LEAD SMELTING AT TARNOWITZ
(September 23, 1905)
The account of the introduction of the Huntington-Heberlein process at Tarnowitz, Prussia, published elsewhere in this issue, is of peculiar interest inasmuch as it tells of the complete displacement by the new process of one of the old processes of lead smelting which had become classic in the art. The roast-reaction process of lead smelting, especially as carried out in reverberatory furnaces, has been for a long time decadent, even in Europe. Tarnowitz was one of the places where it survived most vigorously.
Outside of Europe, this process never found any generally extensive application. It was tried in the Joplin district, and elsewhere in Missouri, with Flintshire furnaces in the seventies. Later it was employed with modified Flintshire and Tarnowitz furnaces at Desloge, in the Flat River district of Missouri, where the plant is still in operation, but on a reduced scale.
The roast-reaction process of smelting, as practised at Tarnowitz, was characterized by a comparatively large charge, slow roasting and low temperature, differing in these respects from the Carinthian and Welsh processes. It was not aimed to extract the maximum proportion of lead in the reverberatory furnace itself, the residue therefrom, which inevitably is high in lead, being subsequently smelted in the blast furnace. Ores too low in lead to be suitable for the reverberatory smelting were sintered in ordinary furnaces and smelted in the blast furnace together with the residue from the other process. In both of these processes the loss of lead was comparatively high. One of the most obvious advantages of the Huntington-Heberlein process is its ability to reduce the loss of lead. The result in that respect at Tarnowitz is clearly stated by Mr. Biernbaum, whose paper will surely attract a good deal of attention.[8]
LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, MO.
By Walter Renton Ingalls
(December 16, 1905)
The roast-reaction method of lead smelting in reverberatory furnaces never found any general employment in the United States, although in connection with the rude air-furnaces it was early introduced in Missouri. The more elaborate Flintshire furnaces were tried at Granby, in the Joplin district, but they were displaced there by Scotch hearths. The most extensive installation of furnaces of the Flintshire type was made at Desloge, in the Flat River district of southeastern Missouri. This continued in full operation until 1903, when the major portion of the plant was closed, it being found more economical to ship the ore elsewhere for smelting. However, two furnaces have been kept in use to work up surplus ore. As a matter of historic interest, it is worth while to record the technical results at Desloge, which have not previously been described in metallurgical literature.
The Desloge plant, which was situated close to the dressing works connected with the mine, and was designed for the smelting of its concentrate, comprised five furnaces. The furnaces were of various constructions. The oldest of them was of the Flintshire type, and had a hearth 10 ft. wide and 14 ft. long. The other furnaces were a combination of the Flintshire and Tarnowitz types. They were built originally like the newer furnaces at Tarnowitz, Upper Silesia, with a rather large rectangular hearth and a lead sump placed at one side of the hearth near the throat end; but good results were not obtained from that construction, wherefore the furnaces were rearranged with the sump at one side, but in the middle of the furnace, as in the Flintshire form. The rectangular shape of the Tarnowitz hearth was, however, retained. Furnaces thus modified had hearths 11 ft. wide and 16 ft. long, except one which had a hearth 13 ft. wide.
The same quantity of ore was put through each of these furnaces, the increase in hearth area being practically of no useful effect, because of inability to attain the requisite temperature in all parts of the larger hearths with the method of heating employed. The men objected especially to a furnace with hearth 13 ft. wide, which it was found difficult to keep in proper condition, and also difficult to handle efficiently. Even the width of 11 ft. was considered too great, and preference was expressed for a 10 ft. width. In this connection, it may be noted that the old furnaces at Tarnowitz were 11 ft. 9 in. long and 10 ft. 10 in. wide, while the new furnaces were 16 ft. long and 8 ft. 10 in. wide (Hofman, “Metallurgy of Lead,” fifth edition, p. 112). All of these dimensions were exceeded at Desloge.
The Flintshire furnaces at Desloge had three working doors per side; the others had four, but only three per side were used, the doors nearest the throat end being kept closed because of insufficient temperature in that part of the furnace. The furnace with hearth 11 × 14 ft. had a grate area of 6.5 × 3 ft. = 19.5 sq. ft.; the 11 × 16 furnaces had grates 8 × 3 = 24 ft. sq. The ratios of grate to hearth area were therefore approximately 1:8 and 1:7.3, respectively. (Compare with ratio of 1:10 at Tarnowitz, and 1:6⅔ at Stiperstones.) The ash pits were open from behind in the customary English fashion. The grate bars were cast iron, 36 in. long. The bars were 1 in. thick at the top, with ⅝ in. spaces between them. The open spaces were 32 in. long, including the rib in the middle. The bars were 4 in. deep at the middle and 2 in. at the ends. The distance from the surface of the grate bars to the fire-door varied in the different furnaces. Some of those with hearths 11 × 16 ft. and grates 8 × 3 ft. had the bars 6 in. below the fire-door; in others the bars were almost on a level with the fire-door.
The furnaces were run with a comparatively thin bed of coal on the grate, and combustion was very imperfect, the percentage of unburned carbon in the ash being commonly high. This was unavoidable with the method of firing employed and the inferior character of the coal (southern Illinois). The excessive consumption of coal was due largely, however, to the practice of raking out the entire bed of coal at the beginning of the operation of “firing down” (beginning the reaction period), when a fresh fire was built with cordwood and large lumps of coal.
Each furnace had two flues at the throat, 16 × 18 in. in size, each flue being provided with a separate damper. Each furnace had an iron chimney approximately 55 ft. high, of which 13 ft. was a brick pedestal (64 × 64 in.) and the remaining 42 ft. sheet steel, guyed. The chimneys were 42 in. in diameter. The distance from the outside end of the furnace to the chimney was approximately 6 ft., and there was consequently but little opportunity for flue dust to collect in the flue. About once a month, however, the chimney was opened at the base and about two wheelbarrows (say 600 lb.) of flue dust, assaying about 50 per cent. lead, was recovered per furnace.
The furnace house was a frame building 45 ft. wide, with boarded sides and a corrugated-iron pitch roof, supported by steel trusses. The furnaces were set in this house, side by side, their longitudinal axes being at right angles to the longitudinal axis of the building. The distance from the outside of the fire-box end of the furnace to the side of the building was 10 ft. The coal was unloaded from a railway track alongside of the building and was wheeled to the furnace in barrows. Some of the furnaces were placed 18 ft. apart; others 22 ft. apart. The men much preferred the greater distance, which made their work easier, an important consideration in this method of smelting.
The hight from the floor to the working door of the furnace was approximately 36 in. The working doors were formed with cast-iron frames, making openings 7 × 11 in. on the inside and 15 × 28 in. on the outside. On the side of the furnace opposite the middle working door was placed a cast-iron hemispherical pot, set partially below the floor-line. This pot was 16 in. deep and 24 in. in diameter; the metal was ¼ in. thick. The distance from the top of the pot to the line of the working door was 31 in.; from the top of the pot to the bottom of the tap-door was 7 in. The tap-door was 4 in. wide and 9 in. high, opening through a cast-iron plate 1½ in. thick. Below the tap-door and on a line with the upper rim of the pot was a tap-hole 3½ in. in diameter. The frames of the working doors had lugs in front, against which the buckstaves bore, to hold the frames in position. All other parts of the sides of the furnace, including the fire-box, were cased with ⅝ in. cast-iron plates, which were obviously too light, being badly cracked.
The cost of a furnace when built in 1893 was approximately $1400, not including the chimney; but with the increased cost of material the present expense would probably be about $2000. Notwithstanding the light construction of the furnaces, repairs were never a large item. Once a month a furnace was idle about 24 hours while the throat was being cleaned out, and every two months some repairing, such as relining the fire-boxes, etc., was required. If repairs had to be made on the inside of the furnace, two days would be lost while it was cooling sufficiently for the men to enter. In refiring a furnace, from 8 to 12 hours was required to raise it to the proper temperature. Out of the 365 days of the year, a furnace would lose from 20 to 25 days, for cleaning the throat and making repairs to the fire-box, arch, etc.
When a furnace was run with two shifts the schedule of operation was as follows:
| Drop charge | 4 a.m. |
| Begin work | 7 a.m. |
| Begin firing down | 11 a.m. |
| Begin first tapping | 1 p.m. |
| Rake out slag | 2.30 p.m. |
| Begin second tapping | 3 p.m. |
| Drop charge | 4 p.m. |
| Begin working | 5.30 p.m. |
| Begin firing down | 11 p.m. |
| Begin first tapping | 1 a.m. |
| Rake out slag | 2.30 a.m. |
| Begin second tapping | 3 p.m. |
With three shifts on a furnace, the schedule was as follows:
| Drop charge | 7 a.m. |
| Begin firing down | 12 a.m. |
| Begin tapping | 1 p.m. |
| Rake out slag | 2 p.m. |
| Begin tapping | 2.30 p.m. |
| Drop charge | 3 p.m. |
| Begin firing down | 8 p.m. |
| Begin tapping | 9 p.m. |
| Rake out slag | 10 p.m. |
| Begin tapping | 10.30 p.m. |
| Drop charge | 11.00 p.m. |
| Begin firing down | 4 a.m. |
| Begin tapping | 5 a.m. |
| Rake out slag | 6 a.m. |
| Begin tapping | 6.30 a.m. |
The hearths were composed of about 8 in. of gray slag beaten down solidly on a basin of brick, which rested on a filling of clay, rammed solid. The hearth was patched if necessary after the drawing of each charge.
The system of smelting was analogous to that which was practiced in Wales rather than to the Silesian, the charges being worked off quickly, and with the aim of making a high extraction of lead directly and a gray slag of comparatively low content in lead. The average furnace charge was 3500 lb. At the beginning of the reaction period about 85 to 100 lb. of crushed fluorspar was thrown into the furnace and mixed well with the charge. The furnace doors were then closed tightly and the temperature raised, the grate having previously been cleaned. At the first tapping about 1200 lb. of lead would be obtained. A small quantity of chips and bark was thrown into the lead in the kettle, which was then poled for a few minutes, skimmed, and ladled into molds, the pigs weighing 80 lb. The skimmings and dross were put back into the furnace. The pig lead was sold as “ordinary soft Missouri.” The gray slag was raked out of the furnace, at the end of the operation, into a barrow, by which it was wheeled to a pile outside of the building. Shipments of the slag were made to other smelters from time to time, 95 per cent. of its lead content being paid for when its assay was over 40 per cent., and 90 per cent. when lower.
Each furnace was manned by one smelter ($1.75) and one helper ($1.55) per shift, when two shifts per 24 hours were run. They had to get their own coal, ore and flux, and wheel away their gray slag and ashes. In winter, when three shifts were run, the men were paid only $1.65 and $1.50 respectively. There was a foreman on the day shift, but none at night. The total coal consumption was ordinarily about 0.8 to 0.9 per ton of ore. Run-of-mine coal was used, which cost about $2 per ton delivered. The coal was of inferior quality, and it was wastefully burned, as previously referred to, wherefore the consumption was high in comparison with the average at Tarnowitz, where it used to be about 0.5 per ton of ore.
The chief features of the practice at Desloge are compared with those at Tarnowitz, Silesia and Holywell (Flintshire), and Stiperstones (Shropshire), Wales, in the following table, the data for Silesia and Wales being taken from Hofman’s “Metallurgy of Lead,” fifth edition, pp. 112, 113.
| Detail | Holywell | Stiper-stones | Tarnowitz | Tarnowitz | Desloge |
|---|---|---|---|---|---|
| Hearth length, ft. | 12.00 | 9.75 | 11.75 | 16.00 | 16.00 |
| Hearth width, ft. | 9.50 | 9.50 | 10.83 | 8.83 | 11.00 |
| Grate length, ft. | 4.50 | 4.50 | 8.00 | 8.00 | 8.00 |
| Grate width, ft. | 2.50 | 2.50 | 1.67 | 1.67 | 3.00 |
| Grate area: hearth area | 1:8 | 1:6⅔ | 1:10 | 1:10 | 1:7-1/3 |
| Charges per 24 hr., | 3 | 3 | 2 | 2 | 3 |
| Ore smelted per 24 hr., lb. | 7,050 | 7,050 | 8,800 | 16,500 | 10,500 |
| Assay of ore, % Pb | 75-80 | 77.5 | 70-74 | 70-74 | 70 |
| Gray slag, % of charge | 12 | 15 | 30 | 27 | |
| Gray slag, % Pb | 55 | 38.8 | 56 | 38 | |
| Men per 24 hr. | 6 | 4 | 4 | 6 | 6 |
| Coal used per ton ore | 0.57-0.76 | 0.56 | 0.46 | 0.50 | 0.90 |
The regular furnace charge at Desloge was 3500 lb. The working of three charges per 24 hours gave a daily capacity of 10,500 lb. per furnace. These figures refer to the wet weight of the concentrate, which was smelted just as delivered from the mill. Its size was 9 mm. and finer. Assuming its average moisture content to be 5 per cent., the daily capacity per furnace was about 10,000 lb. (5 tons) of dry ore.
The metallurgical result is indicated by the figures for two months of operation in 1900. The quantity of ore smelted was 1012 tons, equivalent to approximately 962 tons dry weight. The pig lead produced was 523.3 tons, or 54.4 per cent. of the weight of the ore. The gray slag produced was 262.25 tons, or about 27 per cent. of the weight of the ore. The assay of the ore was approximately 70 per cent. lead, giving a content of 673.4 tons in the ore smelted. The gray slag assayed approximately 38 per cent. lead, giving a content of 99.66 tons. Assuming that 90 per cent. of the lead in the gray slag be recoverable in the subsequent smelting in the blast furnace, or 89.7 tons, the total extraction of lead in the process was 523.3 + 89.7 ÷ 673.4 = 91 per cent. The metallurgical efficiency of the process was, therefore, reasonably high, especially in view of the absence of dust chambers.
The cost of smelting with five furnaces in operation, each treating three charges per day, was approximately as follows:
On the basis of 6.25 tons of wet ore, this would be $4.65 per ton. The actual cost in seven consecutive months of 1900 was as follows: Labor, $1.98 per ton; coal, $1.86; flux and supplies, $0.51; blacksmithing and repairs, $0.39; miscellaneous, $0,017; total, $4.757. If the cost of smelting the gray slag be reckoned at $8 per ton, and the proportion of gray slag be reckoned at 0.25 ton per ton of galena concentrate, the total cost of treatment of the latter comes to about $6.75 per ton of wet charge, or about $7 per ton of dry charge. This cost could be materially reduced in a larger and more perfectly designed plant.
The practice at Desloge did not compare unfavorably, either in respect to metal extracted or in smelting cost, with the roast-reduction method of smelting or the Scotch hearth method, as carried out in the plants of similar capacity and approximately the same date of construction, smelting the same class of ore, but the larger and more recent plants in the vicinity of St. Louis could offer sufficiently better terms to make it advisable to close down the Desloge plant and ship the ore to them. One of the drawbacks of the reverberatory method of smelting was the necessity of shipping away the gray slag, the quantity of that product made in a small plant being insufficient to warrant the operation of an independent shaft furnace.
PART III
SINTERING AND BRIQUETTING
THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN HILL[9]
By E. J. Horwood
(August 22, 1903)
It is well known that, owing to the intimate mixture of the constituents of the Broken Hill sulphide ores, a great deal of crushing and grinding is required to detach the particles of galena from the zinc blende and the gangue; and it will be understood, therefore, that a considerable amount of the material is converted into a slime which consists of minute but well-defined particles of all the constituents of the ore, the relative proportions of which depend on the dual characteristics of hardness and abundance of the various constituents. An analysis of the slime shows the contents to be as follows;
| Galena (PbS) | 24.00 |
| Blende (ZnS) | 29.00 |
| Pyrite (FeS2) | 3.38 |
| Ferric oxide (Fe2O3) | 4.17 |
| Ferrous oxide (FeO) contained in garnets | 1.03 |
| Oxide of manganese (MnO) contained in rhodonite and garnets | 6.66 |
| Alumina (Al2O3) contained in kaolin and garnets | 5.40 |
| Lime (CaO) contained in garnets, etc. | 3.40 |
| Silica (SiO2) | 22.98 |
| Silver (Ag) | .06 |
| 100.48 |
Galena, being the softest of these, is found in the slimes to a larger extent than in the crude ore; it is also, for the same reason, in the finest state of subdivision, as is well illustrated by the fact that the last slime to settle in water is invariably much the richest in lead, while the percentages of the harder constituents, zinc blende and gangue, show a corresponding reduction in quantity, by reason of their being generally in larger sized particles and consequently settling earlier.
The fairly complete liberation of each of the constituent minerals of the ore that takes place in sliming tends, of course, to help the production of a high-grade concentrate by the use of tables and vanners, and undoubtedly a fair recovery of lead is quite possible, even with existing machines, in the treatment of fine slimes; but, owing to the great reduction in the capacity of the machines, which takes place when it is attempted to carry the vanning of the finer slimes too far, and the consequently greatly increased area of the machines that would be necessary, the operation, sooner or later, becomes unprofitable.
The extent to which the vanner treatment of slimes should be carried is, of course, less in the case of those mines owning smelters than with those which have to depend on the sale of concentrates as their sole source of profit. In the case of the Proprietary Company, all slime produced in crushing is passed over the machines after classification. A high recovery of lead in the form of concentrates is, of course, neither expected nor obtained, for reasons already explained; but the finest lead-bearing slimes are allowed to unite with the tailings, which are collected from groups of machines, and are then run into pointed boxes, where, with the aid of hydraulic classification, the fine rich slimes are washed out and carried to settling bins and tanks, where the water is stilled and allowed to deposit its slime, and pass over a wide overflow as clear water. The slime thus recovered amounts to over 1200 tons weekly, or about 11 per cent., by weight, of the ore, and assays about 20 per cent. lead, 17 per cent. zinc, and 18 oz. silver, and represents, in lead value, about 11 per cent. of the original lead contents of the crude ore and rather more than that percentage in silver contents. These slimes are thus a by-product of the mills, and their production is unavoidable; but as they are not chargeable with the cost of milling, they are an asset of considerable value, more especially so since it has been demonstrated that they can be desulphurized sufficiently for smelting purposes by a simple operation, and, at the same time, converted into such a physical condition as renders the material well suited for smelting, owing to its ability to resist pressure in the furnaces.
The Broken Hill Proprietary Company has many thousands of tons of these slimes which the smelters have hitherto been unable to cope with, owing to the roasters being fully occupied with the more valuable concentrates. Moreover, the desulphurization of slimes in Ropp mechanical roasters is objectionable for various reasons, namely, owing to the large amount of dust created with such fine material, resulting injuriously to the men employed; also on account of the reduction in the capacity of the roasters, and consequent increase in working cost, owing to the lightness of the slime, especially when hot, as compared with concentrates, and the necessity for limiting the thickness of material on the bed of the roasters to a certain small maximum. Further, the desulphurization of the slimes is no more complete with the mechanical roasters than in the case of heap roasting, and the combined cost of roasting and briquetting being quite three shillings (or 75c.) per ton in excess of the cost of heap roasting, the latter possesses many advantages. These heaps are being dealt with, preparatory to roasting, by picking down the material in lumps of about 5 in. in thickness, while the fine dry smalls, unavoidably produced, are worked up in a pug mill with water, and dealt with in the same way as the wet slime produced from current work.
The slime, as produced by the mills, is run from bins into railway trucks in a semi-fluid condition, and shortly after being tipped alongside one of the various sidings on the mine is in a fit condition to be cut with shovels into rough bricks, which dry with fair rapidity, and when required for roasting are easily reloaded into railway trucks. As each man can cut about 20 tons of bricks per day, the cost is small. Various other methods of lumping the slime were tried, including trucking the semi-fluid material on movable trams, alongside which were set laths, about 9 in. apart, which enabled long slabs to be formed 9 in. wide and 5 in. thick, which were, after drying, picked up in suitable lumps and loaded in platform trucks, thence on railway trucks. Owing to the inferior roasting that takes place with bricks having flat sides, which are liable to come into close contact in roasting, and to the rather high labor cost, this method was discontinued. Another method was to allow the slime to dry partially after being emptied from railway trucks, and to break it into lumps by means of picks; but this method entailed the making of an increased amount of smalls, besides taking up more siding room, owing to the extra time required for drying, as compared with the method now in use. Ordinary bricking machines could, of course, be used, but when the cost of handling the slime before and after bricking is counted, the cost would be greater than with the simple method now in use; the material being in too fluid a condition for making into bricks until some time elapses for drying, a double handling would be necessitated before sending it to the bricking machine. If, however, the slime could be allowed time to dry sufficiently in the trucks, bricking by machinery would probably be preferable. Rather more than 10 per cent. of smalls is made in handling the lumps in and out of the railway trucks, and this is, as already noted, worked up with water in a pug mill at the sintering works, and used partly for covering the heaps with slime to exclude an excessive amount of air. The balance is thrown out and cut into bricks, as already described.
At the heaps the lumps are at present being thrown from one man to another to reach their destination in the heap, but the sidings have been laid out in duplicate with a view to enabling traveling cranes to be used on the line next the heap, the lumps to be loaded primarily into wooden skips fitting the trucks. It is probable, however, that the lumps will require to be handled out of the skips into their place in the heap, as the brittle nature of the material may be found to render automatic tipping impracticable. A considerable saving in labor would nevertheless accompany the use of cranes, which would likewise be advantageous in loading the sintered material.
In order to reduce the inconvenience arising from fumes, length is very desirable in siding accommodation, so that heap building may be carried on at a sufficient distance from the burning kilns. It is for the same reason preferable to build in a large tonnage at one time, lighting the heaps altogether. As the heaps burn about two weeks only, long intervals intervene, during which the fumes are absent.
In the experimental stages of slime roasting, fuel, chiefly wood, was used in quantities up to 5 per cent., and was placed on the ground at the bottom of the heap, where also a number of flues, loosely built bricks, were placed for the circulation of air. The amount of fuel used has, however, been gradually reduced, until the present practice of placing no fuel whatever in the bottom was arrived at; but instead less than 1 per cent. of wood is now burned in small enlargements of the flues, under the outer portion of the pile, and placed about 12 ft. apart at the centers. This is found to be sufficient to start the roasting operation within 24 hours of lighting, after which no further fuel is necessary.
As regards the dimensions of the heaps, the width found most suitable is 22 ft. at the base, the sides sloping up rather flatter than one to one, with a flat section on top reaching about 7 ft. in hight. As there is always about 6 in. of the outer crust imperfectly roasted, it is advisable to make the length as great as possible, thus minimizing the surface exposed. The company is building heaps up to 2000 ft. long.
During roasting care is required to regulate the air supply, the object being to avoid too fierce a roast, which tends to sinter and partially fuse the material on the outer portions of the lumps, while inside there is raw slime. By extending the roast over a longer period this is avoided, and a more complete desulphurization is effected. Experiments conducted by Mr. Bradford, the chief assayer, demonstrated that, at a temperature of 400 deg. C., the sulphide slime is converted into basic sulphate, while at a temperature of 800 deg. C. the material becomes sintered owing to the decomposition of the basic sulphate and the formation of fusible silicate of lead.
In practice, the sulphur contents of the material, which originally are about 14 per cent., become reduced to from 6.5 to 8.5 per cent., half in the form of basic sulphate and half as sulphides; much of the material sinters and becomes matted together in a fairly solid mass. The heaps are built without chimneys of any kind; a strip about 5 ft. wide along the crest of the pile is left uncovered by plastered slime, and this, together with the open way in which the lumps are built in, allows a natural draft to be set up, which can be regulated by partly closing the open ends of the flues at the base of the pile. Masonry kilns were used in the earlier stages with good results, which, however, were not so much better than those obtained by the heap method as to justify the expense of building, taking into consideration, too, the extra cost of handling the roasted material in the necessarily more confined space.
Much interest has been taken in the chemical reactions which take place in the operation of desulphurization of these slimes, it being contended, on the one hand, that the unexpectedly rapid roast which takes place may be due to the sulphide being in a very fine state of subdivision, and more or less porous, thus allowing the air ready access to the sulphur, producing sulphurous acid gas (SO2). On the other hand, others, of whom Mr. Carmichael is the chief exponent, claim that several reactions take place during the operation, connected with the rhodonite and lime compounds present in the slimes, which he describes as follows:
“The temperature of the kilns having reached a dull red heat, the rhodonite (silicate of manganese) is converted into manganous oxide and silica; at a rather higher temperature the calcium compounds are also split up, with formation of calcium sulphide, the sulphur being provided by the slimes. The air permeating the mass oxidizes the manganese oxide and calcium sulphide into manganese tetroxide and calcium sulphate respectively, as shown as follows;
- 3MnO + O = Mn3O4
- CaS + 4O = CaSO4,
and, as such, are carriers of a form of concentrated oxygen to the sulphide slimes, with a corresponding reduction to manganous oxide and calcium sulphide, as shown by the following equation, in the case of lead:
- PbS + 4Mn3O4 = PbSO4 + 12MnO
- PbS + CaSO4 = PbSO4 + CaS.
The oxidation of the manganous oxide and calcium sulphide is repeated, and these alternate reactions recur until the desulphurization ceases, or the kiln cools down to a temperature below which oxidation cannot occur. These reactions, being heat-producing, provide part of the heat necessary for desulphurization, which is brought about by certain concurrent reactions between metallic sulphates and sulphide.
“The first that probably occurs is that in which two equivalents of the metallic sulphide react on one of the metallic sulphate with reduction to the metal, metallic sulphide, and sulphurous acid, as shown by the following equation in the case of lead:
- 2PbS + PbSO4 = 2Pb + PbS + 2SO2.
“The metal so formed, in the presence of air, is oxidized, and in this state reacts on a further portion of the metallic sulphide produced, with an increased formation of metal and evolution of sulphurous acid, according to the following equation, in the case of lead:
- 2PbO + PbS = Pb + SO2.
“The metal so produced in this reaction is wholly reoxidized by the oxygen of the air current, and being free to react on still further portions of the metallic sulphide, repeats the reaction, and becomes an important factor in the desulphurizing of the undecomposed portion of the material. As the desulphurization proceeds, and the sulphate of metal accumulates, reactions are set up between the metallic sulphide and different multiple proportions of the metallic sulphate, with the formation of metal, metallic oxide, and evolution of sulphurous acid, as follows:
“With two equivalents of metallic sulphate to one equivalent of metallic sulphide, in the case of lead, according to the following equation:
- PbS + 2PbSO4 = 2PbO + Pb + 3SO2.
“With three equivalents of metallic sulphate to one of metallic sulphide, in the case of lead, according to the following equation:
- PbS + 3PbSO4 = 4PbO + 4SO2.”
The volatility of sulphide of lead—especially in the presence of an inert gas such as sulphurous acid—being greater than that of the sulphate, oxide, or the metal itself, it might be thought that the conditions are conducive to a serious loss of lead. This, however, is reduced to a minimum, owing to the easily volatilized sulphide being trapped, as non-volatile sulphate, by small portions of sulphuric anhydride (SO3), which is formed by a catalytic reaction set up between the hot ore, sulphurous acid, and the air passing through the mass. Owing to the non-volatility of the silver compounds in the slimes, the loss of this metal has been found to be inappreciable. The zinc contents of the slime are reduced appreciably, thus rendering the material more suitable for smelting. After desulphurization ceases, a few days are allowed for cooling off. On the breaking up of the mass for despatch to the smelters, as much of the lower portion of the walls is left intact as possible, so that it can be utilized for the next roast, thus avoiding the re-building of the whole of the walls.[10]
THE PREPARATION OF FINE MATERIAL FOR SMELTING
By T. J. Greenway
(January 12, 1905)
In the course of smelting, at the works of the company known as the Broken Hill Proprietary Block 14, material which consisted chiefly of silver-lead concentrate and slime, resulting from the concentration of the Broken Hill complex sulphide ore, I had to contend with all the troubles which attend the treatment of large quantities of finely divided material in blast furnaces. With the view of avoiding these troubles, I experimented with various briquetting processes; and, after a number of more or less unsatisfactory experiences, I adopted a procedure similar to that followed in manufacturing ordinary bricks by what is known as the semi-dry brick-pressing process. This method of briquetting not only converts the finely divided material cheaply and effectively into hard semi-fused lumps, which are especially suitable for the heavy furnace burdens required by modern smelting practice, but also eliminates sulphur, arsenic, etc., to a great extent; therefore, it is capable of wide application in dealing with concentrate, slime, and other finely divided material containing lead, copper and the precious metals.
This briquetting process comprises the following series of operations:
1. Mixing the finely divided material with water and newly slaked lime.
2. Pressing the mixture into blocks of the size and shape of ordinary bricks.
3. Stacking the briquettes in suitably covered kilns.
4. Burning the briquettes, so as to harden them, without melting, at the same time eliminating sulphur, arsenic, etc.
1. The material is dumped into a mixing plant, together with such proportions of screened slaked lime (usually from three to five per cent.) and water as shall produce a powdery mixture which will, on being squeezed in the hand, cohere into dry lumps. In preparing the mixture, it is well to mix sandy material with suitable proportions of fine, such as slime, in order that the finer material may act as a binding agent.
The mixer used by me consists of an iron trough, about 8 ft. long, traversed by a pair of revolving shafts, carrying a series of knives arranged screw-fashion; and so placed that the knives on one shaft travel through the spaces between the knives on the other shaft. The various materials are dumped into one end of the mixing trough, from barrows or trucks, and are delivered continuously at the other end of the trough, into an elevator which conveys the mixture to the brick-pressing plant.
2. The plant employed was the semi-dry brick-press. This machine receives the mixture from the elevators, and delivers it in the form of briquettes, which can at once be stacked in the kilns. It was found that such material as concentrate and slime has comparatively little mobility in the dies during the pressing operation; this necessitates the use of a device which provides for the accurate filling of the dies. It was also found that the materials treated by smelters vary in compressibility, and this renders necessary the adoption of a brick-pressing plant having plungers which are forced into the dies by means of adjustable springs, brick-presses having plungers actuated by rigid mechanism being extremely liable to jam and break.
3. Briquettes made from such material as concentrate and slime vary in fusibility; they are also combustible, and while being burned they produce large quantities of smoke containing sulphurous acid and other objectionable fumes. It is therefore necessary that such briquettes be burned in kilns provided with arrangements for accurately controlling the burning operations, and for conveniently disposing of the smoke. Suitable kilns, which will contain from 30 to 50 tons of briquettes per setting, are employed for this purpose. Regenerative kilns of the Hoffman type might be used for dealing with some classes of material, but, for general purposes, the kilns as designed here will be found more convenient.
The briquettes are stacked according to the character of the material and the object to be obtained. The various methods of stacking, and the reasons for adopting them, can be readily learned by studying ordinary brick-burning operations in any large brick-yard. After the stacking is complete the kiln-fronts are built up with burnt briquettes produced in conducting previous operations, and all the joints are well luted.
4. In burning briquettes made from pyrite or other self-burning material, it is simply necessary to maintain a fire in the kiln fireplaces for a period of from 10 to 20 hours. When it is judged that this firing has been continued long enough, the fire-bars are drawn and the fronts are luted with burnt briquettes in the same manner as the kiln-fronts. Holes about two inches square are then made in these lutings, through which the air required for the further burning of the briquettes is allowed to enter the kilns under proper control. After the fireplaces are thus closed the progress of the burning, which continues for periods of from three to six days, is watched through small inspection holes made in the kiln-fronts; and when it is seen that the burning is complete the fronts are partially torn away, in order to accelerate the cooling of the burnt briquettes, which are broken down and conveyed to the smelters as soon as they can be conveniently handled.
When briquettes made from pyrite concentrate, or of other free-burning material, are thus treated, they are not only sintered but they are also more or less effectively roasted, and it may be taken for granted that any ore which can be effectively roasted in the lump form in kilns or stalls will form briquettes that will both sinter and roast well; indeed, one may say more than this, for briquettes which will sinter and roast well can be made from many classes of ore that cannot be effectively treated by ordinary kiln-and stall-roasting operations; and, moreover, good-burning briquettes may be made from mixtures of free-burning and poor-burning material. Briquettes containing large proportions of pyrite or other free-burning material will, unless the air-supply is properly controlled, often heat up to such an extent as to fuse into solid masses, much in the same manner as matte of pyritic ore will melt when it is unskilfully handled in roasting. In dealing with material which will not burn freely, such as roasted concentrate, the briquetting is conducted with the intention of sintering the material; and in this case the firing of the kilns is continued for periods of from three to four days, the procedure being similar in every way to that followed in burning ordinary bricks.
When conducting my earlier briquetting operations I made the briquettes by simply pugging the finely divided material, following a practice similar to that adopted in producing “slop-made” bricks by hand. This method of making the briquettes was attended with a number of obvious disadvantages, and was abandoned as soon as the semi-dry brick-pressing plant became available. The extent to which this process, or modifications of it, may be applied is shown by the fact that, following upon information given by me, the Broken Hill Proprietary Company adopted a similar method of sintering and roasting slime, consisting of about 20 per cent. galena, 20 per cent. blende, and 60 per cent. silicious gangue. The procedure followed in this case consisted of simply pugging the slime, and running the pug upon a floor to dry; afterward cutting the dried material into lumps by means of suitable cutting tools, and then piling the lumps over firing foundations, following a practice similar to that pursued in conducting ordinary heap-roasting. This company is now treating from 500 to 1000 tons of slime weekly in this manner. It is, however, certain that better results would attend the treatment of this material by making this slime into briquettes and burning them in kilns.
The cost of briquetting and burning material in the manner first described, with labor at 25c. per hour, and wood or coal at $4 per ton, amounts to from $1 to $1.50 per ton of material.
THE BRIQUETTING OF MINERALS
By Robert Schorr
(November 22, 1902)
The value of briquetting in connection with metallurgical processes and the manufacture of artificial stone is well understood and appreciated. In smelting plants there is always more or less flue dust, fine ores, and sometimes fine concentrates to be treated, but the charging of such fine material directly into a furnace would cause trouble and irregularities, and would lessen its capacity also. As mineral briquetting cannot be effected without considerable wear upon the machinery and without quite appreciable expense in binder, labor, and handling, many smelters try to avoid it.
The financial question, however, is not as serious as it may at first appear, and taking the large output of modern briquetting machines in consideration, the cost for repairs amounts only to a few cents per ton of briquetted material. The total cost depends in the first place on the cost of labor, power and the binder, and in most American smelters it varies between $0.65 and $1.25 per ton of briquettes.
Ordinary brick presses, with clay as a binder, were used in Europe as well as in this country, but they are too slow and expensive for large propositions and the presence of clay is usually undesirable.
The English Yeadon (fuel) press has also been used for some years at the Carlton Iron Company’s Works at Ferryhill in England, and at the Ore and Fuel Company’s plant at Coatbridge in the same country; also by some Continental firms. Dupuis & Sons, Paris, furnished a few presses which are mostly used for manganese and iron ores and pyrites. In some localities coke dust is added. The making of clay briquettes or mud-cakes is the crudest form of briquetting; but while heat has to be expended to evaporate the 40 to 50 per cent. of moisture in them, and while considerable flue dust is made, this method is better than feeding fine ore or flue dust directly into the furnace.
The only other method of avoiding briquetting is by fusing ore fines in slagging reverberatory furnaces and by adding flue dust in the slagging pit, thus incorporating it with the slagging ore. This is practised sometimes in silver-lead smelters, but in connection with copper or iron smelters it is not practicable.
In briquetting minerals a thorough mixing and kneading is of the first importance. If this is done properly a comparatively low pressure will suffice to create a good and solid briquette, which after six to eight hours of air-drying, or after a speedier elimination of the surplus of moisture in hot-air chambers, will be ready for the furnace charge. A good briquette should permit transportation without excessive breakage or dust a few hours after being made, and it should retain its shape in the furnace until completely fused, so as to create as little flue dust as possible. The briquette should be dense, otherwise it will crumble under the influence of bad weather.
The two presses on the American machinery market are the type built by the Chisholm, Boyd & White Company, of Chicago, and the briquetting machine manufactured by the H. S. Mould Company, of Pittsburg. Both are extensively used, and in many metallurgical plants it will pay well to adopt them.
From 4 to 6 per cent. of milk of lime is generally used as binder, and this has a desirable fluxing influence also. A complete outfit comprises, besides the press, a mixer for slacking the lime, and a feed-pump which discharges the liquid in proportion into the main mixer wherein the ore fines, flue dust, or concentrates are shoveled.
The Chisholm, Boyd & White Company’s press makes 80 briquettes per minute, which, with a new disk, are of 4 in. diameter and 2½ in. hight, thus giving about 872 cu. ft. of briquette volume per 10 hours, or 50 to 80 tons, depending on the weight of the material. With the wear of the disk the hight of the briquettes is reduced and consequently the capacity of the machine also. The disk weighs about 1600 lb., and as most large smelters have their own foundries it can be replaced with little expense. About 30 effective horse-power is usually provided for driving the apparatus. The machine is too well known to metallurgists and engineers to require further comment or description.
The H. S. Mould Company has also succeeded in making its machine a thorough practical success. This machine is a plunger-type press. The largest press built employs six plungers, and at 25 revolutions it makes 150 briquettes of 3 in. diameter and 3 in. hight, or 1080 cu. ft. per 10 hours. Its rated capacity is 100 tons per 10 hours.
In using a plunger-type press the material should not contain more than 7 per cent. mechanical moisture. If wet concentrates have to be briquetted it is necessary to add dry ore fines or flue dust to arrive at a proper consistency. The briquettes are very solid and only air-drying for a few hours is necessary.
The cylindrical shape of briquettes is very good, as it insures a proper air circulation in the furnace and consequently a rapid oxidation and fusion.
The wear of the Mould Company’s press is mostly confined to the chilled iron bushings and to the pistons. Auxiliary machinery consists of the slacker, the feeder and the main mixer. The press is of a very substantial design, and it is claimed that the cost of repairs does not amount to more than 3c. per ton of briquettes.
Wear and tear is unavoidable in a crude operation like briquetting; to treat flue dust, ore fines, and fine concentrates successfully, it is almost absolutely necessary to resort to it.
Edison used a number of intermittent-acting presses at his magnetic iron-separation works in New Jersey, but this plant shut down some time ago.
A BRICKING PLANT FOR FLUE DUST AND FINE ORES
By James C. Bennett
(September 15, 1904)
The plant, which is here described, for bricking fine ores and flue dust, was designed and the plans produced in the engineering department of the Selby smelter. The machinery contained in the plant consists of a Boyd four-mold brick press, a 7 ft. wet pan or Chile mill, a 50 h.p. induction motor, and a conveyor-elevator, together with the necessary pulleys and shafting.
The press, Chile mill, and motor need no special mention, as they all are from standard patterns and bought, without alterations, from the respective builders. The Chile mill was purchased from the builders of the brick press. The conveyor-elevator was built on the premises and consists of a 14 in. eight-ply rubber belt, with buckets of sheet steel placed at intervals of 6 in., running over flanged pulleys. The buckets, or more properly speaking the flights, are made from No. 12 steel plate, flanged to produce the back and ends, with the ends secured to the flanged bottom by one rivet in each. The plant has been in operation for sixteen months and there have been few or no repairs to the elevator, except to renew the belt, which is attacked by the acid contained in the charges. This first belt was in continuous use for nine months. As originally designed, the capacity was 100 tons per day of 12 hours, but this was found to require a speed so high that the workmen were unable to handle the output of the press. The speed was, consequently, reduced about 25 per cent., which brings the output down to about 75 tons per day. This output, as expressed in weight, naturally varies somewhat owing to the variation in the weight of the material handled.
It is probable that the capacity could be increased to about 90 tons by enlarging the bricks, which could be done, but would require a considerable amount of alteration in the machine, as it is designed to produce a standard sized building brick. By this method of increase, however, the work of handling would not be materially increased, because the number of bricks would be the same as with the present output of 75 tons; there would be about 16 per cent. more to handle, by weight. Working on the basis of 100 tons capacity, the bins were designed to afford storage room for about three days’ run, or a little over 300 tons. The bins are made entirely of steel, in order that the hot material may be dumped into them directly from the roasting furnaces, thus saving one handling. In order that there may be room for several kinds of material, the bins are divided into seven compartments, three on one side and four on the other. The lower part is of ⅜ in. steel plate, and the upper, about one-half the hight, of 5/16 in. plate.
It may be well to call attention to the method of handling the material, preparatory to its delivery to the brick press. The bins are constructed, as will be seen by the drawing, with their floor set 2.5 ft. above the working floor, which enables the workmen to reach the material with a minimum effort. The floor of the bins project 2.5 ft. in front of the face, thus forming a platform on which the shoveling may be done without the necessity of bending over. In this projecting platform are cut rectangular holes 12 × 18 in., which are placed midway between the openings in the front of the bins and furnished with screens to stop any stray bolts or other coarse material that might injure the press. This position of the holes through the platform was adopted so that, in the event of the material running out beyond the opening in the face, it would not fall directly upon the floor. Two buckets are provided, with a capacity of 7 cu. ft. each, which is the size of a single charge of the Chile mill. These buckets have a hopper-shaped bottom fixed with a swinging gate which is operated by the foot; thus the bucket can be run over the pan of the Chile mill and the charge dumped directly into it. The buckets run on an overhead iron track (1 in. by 3 in.) hung 7 ft. in the clear, above the floor.
The method of making up the charge is as follows: The bucket is run under the hole in the platform nearest to the compartment containing the material of which the charge is partly composed, and a predetermined number of shovelfuls is drawn out and put into the bucket, which is then pushed on to the next compartment from which material is wanted, where the operation is repeated. After charging into the bucket the requisite amount of ore or flue dust, the bucket is run to the back of the building, where the necessary amount of lime (slaked) is added. By putting the lime in last, it is so surrounded by the dust or ore that it has not the opportunity to stick to the sides of the bucket in discharging, as it otherwise would.
Fig. 1 (a).—Plant for Bricking Ores, Selby Smelter. (Plan.)
The number of men required to operate the entire plant, exclusive of those employed in bringing the material to the bins and emptying the cars into them, is 12, placed as follows; One preparing the lime for use, one removing the charge from the mill and supplying the elevator-conveyor, which is accomplished by means of a specially shaped, long-handled shovel; one keeping the supply spout of the press clear (an attempt was made to do this mechanically, but was found to be unsuccessful, owing to the extremely sticky nature of the material, and so was discarded in favor of manual labor); one to control the press in case of mishap and to keep the dies clean; one oiler; three receiving the bricks from the press and taking the brick-loaded cars from the press to the drying-house, and two placing the bricks on the shelves.
Fig. 1 (b).—Plant for Bricking Ores, Selby Smelter. (Elevation.)
The drying-house scarcely requires description; it is but a roofed shed, without sides, fitted with stalls into which the bricks are set on portable shelves, as close as working conditions will permit. The means of drying, at the present time, is by the natural circulation of air, but a mechanical system is in contemplation, by which the air will be drawn into the building from the outside and forced to find its way out through the bricks. The drying-house is adjacent to the pressing plant, in fact forms the back of it, so that there is a minimum distance to haul the product. The time required for drying the bricks sufficiently for them to withstand the necessary handling is, depending on the weather, from two to eight days, the usual time being about three days.
PART IV
SMELTING IN THE BLAST FURNACE
MODERN SILVER-LEAD SMELTING[11]
By Arthur S. Dwight
(January 10, 1903)
The rectangular silver-lead blast furnace developed in the Rocky Mountains has an area of 42 × 120 to 48 × 160 in. at the tuyeres; 54 × 132 to 84 × 200 in. at the top; and hight from tuyere level to top of charge of 15 to 21 ft. Such a furnace smelts 80 to 200 tons of charge (ore and flux, but not slag and coke) per 24 hours. The slag that has to be resmelted amounts to 20 to 60 per cent. of the charge. Coke consumption is 12 to 16 per cent. of the charge. The blast pressure ranges from 1.5 to 4 lb. per square inch, averaging close to 2 lb. Gases of hand-charged furnaces are taken off through an opening below the charge-floor, the furnace being fed through a slot (about 20 in. wide, extending nearly the whole length of the furnace) in the iron floor-plates; or through a hood (brick or sheet iron) above the charge-floor level, with a down-take to the flues, charge-doors being provided on each side of the hood, extending preferably the whole length of the furnace and usually having a sill a few inches high which compels the feeder to lift his shovel.
When a silver-lead blast furnace is operating satisfactorily, the following conditions should obtain; (1) A large proportion of the lead in the charge should appear as direct bullion-product at the lead-well. (2) The slag should be fluid and clean. (3) The matte should be low in lead. (4) The furnace should be cool and quiet on top, making a minimum quantity of lead-fume and flue-dust, and the charges should descend uniformly over the whole area of the shaft. (5) The furnace speed should be good. (6) The furnace should be free from serious accretions and crusts; that is to say, the tuyeres should be reasonably bright and open, and the level of the lead in the lead-well should respond promptly to variations of pressure, caused by the blast and by the hight of the column of molten slag and matte inside the furnace—an indication that ample connection exists between the smelting column and the crucible. Good reduction (using that term to express the degree in which the furnace is manifesting its reducing action) is obtained when the first three of the above conditions are satisfied.
For any given furnace there are five prime factors, the resultant of which determines the reduction, namely: (a) Chemical composition of the furnace charges; (b) proportion and character of fuel; (c) air-volume and pressure, to which might perhaps also be added temperature of blast; for, although hot blast has not yet been successfully applied in lead-smelting practice, I believe it is only a question of time when it will be; (d) dimensions and proportions of smelting furnace; (e) mechanical character and arrangement of the smelting column.
All but one of the above factors can be intelligently gaged. The mechanical factor, however, can be expressed only in generalities and indefinite terms. A wise selection of ores and proper preliminary preparation, crushing the coarse and briquetting the fine, will do much to regulate it, but all this care may be largely nullified by careless feeding. The importance and possibilities of the mechanical factor are generally overlooked and its symptoms are wrongly diagnosed. For instance, the importance of slag-types has undoubtedly been considerably exaggerated at the expense of the mechanical factor. Slags seldom come down exactly as figured. We must know our ores and apply certain empirical corrections to the iron, sulphur, etc., based on previous experience with the ores; but these empirical corrections may represent also an unformulated expression of the influence of the mechanical factor on the reduction—a function, therefore, of the ruling physical complexion of the ores, and the peculiarities of the feeding habitually maintained in the works concerned. With a given ore-charge large reciprocal variations may be produced in the composition of slag and matte by merely changing the mechanical conditions of the smelting column, and since the efficient utilization of both fuel and blast must be controlled in the same way, the mechanical factor may be considered, perhaps, the dominating agent of reduction. Inasmuch as there is no way of gaging it, however, the only recourse is to seek a correct adjustment and maintain it as a positive constant, after which slag, fuel and blast may be with much greater certainty adjusted toward efficiency of furnace work and metal-saving.
Behavior of Iron.—The output of lead is so dependent upon the reactions of the iron in the charge that the chief attention may well be fixed upon that metal as the key to the situation. The success of the process depends largely upon reducing just the right amount of iron to throw the lead out of the matte, the remainder of the iron being reduced only to ferrous oxide and entering the slag. Too much iron reduced will form a sow in the hearth. Iron is reduced from its oxides principally by contact with solid incandescent carbon, and by the action of hot carbon monoxide. Reduction by solid carbon is the more wasteful, but there is in lead smelting an even more serious objection to permitting the reduction to be accomplished by that means, which leads to comparatively hot top and more or less volatilization of lead. Reduction by carbon monoxide is the ideal condition for the lead furnace. It means keeping the zone of incandescence low in the charge column, leaving plenty of room above for the gases to yield up their heat to, and exercise their reducing power on, the descending charge, so that by the time they escape they will be well-nigh spent. Their volume and temperature will be diminished, and the low velocity of their exit will tend to minimize the loss of lead in fume and flue dust.
The idea that high temperatures in lead blast furnaces should be avoided is based on a misconception. Temperatures must exist which are sufficiently high to volatilize all the lead in the charge, if other conditions permit. A high temperature before the tuyeres means fast smelting; and fast smelting, under proper conditions, means a shortening of the time during which the lead is subject to scorifying and volatilizing influences. A rapidly descending charge, constantly replenished with cold ore from above, absorbs effectively the heat of the gases and acts as a most efficient dust and fume collector. In considering long flues, bag-houses, etc., it should be kept in mind that the most effective dust collector ought to be the furnace itself.
In the practice of twelve years ago and earlier, particularly when using mixed coke and charcoal, reduction by carbon was probably the rule; and the percentage of fuel required was very high. There is good reason to think we have still much room for improvement along this line in our average practice of today.
Volume of Blast.—It is customary to supply a battery of furnaces from a large blast main, connected with a number of blowers. Inasmuch as the air will take preferably the line of least resistance, if the internal resistance of any one furnace be increased the volume of air it will take will be diminished and the others will be favored unduly. Only by keeping all the furnaces on approximately the same charge, with the same hight of smelting column, can anything like uniformity of operation and close regulation be secured. The rational plan would seem to be to have a separate blower, of variable speed, directly connected to each furnace, but this plan, which has had a number of trials, has usually been abandoned in favor of the common blast main. Trials by myself, extending over considerable periods, have been so uniformly favorable, however, that I am forced to ascribe the failure of others to some outside reason.
The peculiar atmosphere required in the lead blast furnace depends upon the correct proportion of two counteractive elements, carbon and oxygen. If given too much air the furnace will show signs of deficient reduction, commonly interpreted as calling for more fuel, which will be sheer waste since its object is to burn up surplus air. There will be an additional waste through the extra coal burned under the steam boilers. The true remedy would be to cut down the quantity of air. Burning up excessive coke is as hard work as smelting ore. Too much fuel invariably slows up a furnace; it also drives the fire upward and gives predominance to reduction by solid carbon. The maintenance of a minimum fuel percentage, with a correctly adjusted volume of air, will tend to promote the conditions under which iron will be reduced by the gases, rather than by solid carbon.
Pressure of Blast.—Pressure necessarily involves resistance; and the blast-pressure, as registered by a simple mercury-gage on the bustle-pipe, may be increased in two ways: (1) By increasing the volume of air forced through the interstices in the charge. This is the wrong way; but, unfortunately, it is only too common in our practice, and therefore deserves to be mentioned, if only to be condemned. (2) By leaving the volume of air unchanged, but increasing the friction offered by the interstitial channels, either by making them smaller in aggregate cross-section (which means a finer charge), or by making them longer (which means a higher smelting column). A correctly graduated internal resistance is, therefore, the only true basis for a high blast furnace, which, when so produced, will bring about rapid smelting, a low zone of incandescence, and a very vigorous action upon the ores by the gases in their retarded ascent through the charge column. These conditions promote the reduction of iron by CO. The adjustment of internal resistance, which is thus clearly the main factor, can be accomplished only by the correct feeding of the furnace.
Feeding the Charge.—It is self-evident that, the more thorough the preliminary preparation of the charge before it reaches the zone of fusion, the more rapidly can the actual smelting proceed. A piece of raw ore that finds itself prematurely at the tuyeres, without having been subjected to the usual preparatory processes of drying, heating, reduction, etc., must remain there until it is gradually dissolved or carried away mechanically in the slag. Any such occurrence must greatly retard the process. It would seem, by the same reasoning, that an intimate mixture of the ingredients of the charge should expedite the smelting, and I advocate the intimate mixture of the charge ingredients in all cases.
The theory of feeding is simple, but not so the practice. If the charge column were composed of pieces of uniform size, the ascending gases would find the channel of least resistance close to the furnace walls and would take it preferably to the center of the shaft. The more restricted channel would necessitate a higher velocity, so that not only would the center of the charge be deprived of the action of the gases, but also the portion traversed would be overheated; many particles of ore would be sintered to the walls or carried off as flue dust; slag would form prematurely; fuel would be wasted; in short, all the irregularities and losses which accompany over-fire would be experienced. In practice the charge is never uniform, but is a mixture of coarse and fine. By lodging the finer material close to the walls and placing the coarser in the center, an adjustment may be made which will cause the gases to ascend uniformly through the smelting column. A furnace top smoking quietly and uniformly over its whole area is the visible sign of a properly fed furnace.
Effect of Large Charges.—It has frequently been remarked that, within certain limits, large charges give more favorable results than small ones; and numerous attempts have been made to account for this fact. My observations lead me to offer the following as a rational explanation—at least in cases where ore and fuel are charged in alternate layers. Large ore-charges mean correspondingly large fuel-charges. The gases can pass readily through the coke; and hence each fuel-zone tends to equalize the gas currents by giving them another opportunity to distribute themselves over the whole furnace area, while each layer of ore subsequently encountered will blanket the gases, and compel them to force a passage under pressure, which is the manner most favorable to effective chemical action.
In mechanically fed furnaces the charges of ore and fuel are usually dropped in simultaneously from a car and the separate layers thus obliterated, and the distributing zones which are such a safeguard against the consequences of bad feeding are lacking, hence more care must be exercised to secure proper placing of the coarse and fine material. This may throw some light on the failure of most of the early attempts at mechanical feeding.
Mechanical Character of Charge.—Very fine charges blanket the gases excessively and cause them to break through at a few points, forming blow-holes, which seriously disturb the operation, cause loss of raw ore in the slag, and are accompanied by all the evils of over-fire. A charge containing a few massive pieces, the rest being fine, is a still more unfavorable combination. A very coarse charge permits too ready an exit to the gases, and in the end tends likewise to over-fire and poor reduction. The remedy is to briquette the fine ore (though preferably not all of it), and crush the coarse to such degree as to approach an ideal result, which may be roughly described as a mixture in which about one-third is composed of pieces of 5 to 2 in. in diameter, one-third pieces of 2 to 0.5 in., and the remaining third from 0.5 in. down. The coke is better for being somewhat broken up before charging, and a reasonable amount of coke fines, such as usually accompanies a good quality of coke, is not in the least detrimental. The common practice of handling the coke by forks and throwing away the fines is to be condemned as an unwarranted waste of good fuel. The slag on the charge should be broken to pieces at most 6 in. in diameter. The common practice of throwing in whole butts of slag-shells is bad. There is no economy in using the slag hot; cold charges, not hot, are what we want. A reasonable amount of moisture in the charge is beneficial, providing it be in such form as to be readily dried out. It is often advantageous to wet the ore mixtures while bedding them, or to sprinkle the charges before feeding. The driving off of this water must consume fuel, but not so much as if the smelting zone crept up. Large doses of water applied directly to the furnace are unpardonable under any circumstances, however, though they are sometimes indulged in as a drastic measure to subdue excessive over-fire when other and surer means are not recognized. One of the chief merits of moderate sprinkling before charging is that it gives in many cases a more favorable mechanical character, approximating a lumpy condition in too fine a charge, and assisting to pack a too coarse one.
Different Behavior of Coarse and Fine Ore.—In taking up a shovelful of ore, the fine will be observed to predominate in the bottom and center, and the coarse on the top and sides. When thrown from the shovel, the coarse will outstrip the fine and fall beyond it. In making a conical pile the coarse ore will roll to the base, leaving the fine near the apex. This difference in the action of the mobile coarse ore and the sluggish fines is the key to the practical side of feeding, both manual and mechanical. It is not sufficient to tell the feeder to throw the coarse in the middle and the fine against the sides; if it be easier to do it some other way such instructions will count for little. The desired result can be best secured by making the right way easier than the wrong way.
It is generally conceded that the open-top furnaces, fed by hand through a slot in the floor-plates, do not give as satisfactory results as the hooded furnaces with long feed-doors on both sides. In the open-top furnace it is comparatively difficult to throw to the sides; the narrower the slot the greater the difficulty. The major part of the charge will drop near the center, making that place higher than the sides. The fine ore will tend to stay where it falls, while the coarse will tend to roll to the sides, thus leading to an arrangement of the charge just the reverse of what it ought to be. In the hooded furnace most of the material will naturally fall near the doors, causing the sides to be higher than the center toward which the coarse will roll, while the force of the throw as the ore is shoveled in will also have a tendency to concentrate the coarse material in the center.
Once a proper balance of conditions has been found, absolute regularity of routine is the secret of good results. An experienced and intelligent feeder owes his merit to his conscientious regularity of work. He may have to vary his program somewhat when he encounters a furnace that is suffering from the results of bad feeding by a predecessor; but his guiding principle is first to restore regularity, and then maintain it. A poor feeder can bring about, in a single shift, disorders that will require many days to correct, if indeed they are corrected at all during the campaign. The personal element is productive of more harm than good.
Mechanical Feeding.—If it be admitted that the work of a feeder is the better the more it approximates the regularity of that of a machine, it ought to be desirable to eliminate the personal factor entirely and design a machine for the purpose, which would be a comparatively simple matter if it be known just what we want to accomplish. No valid ground now exists for prejudice against mechanical feeding in lead smelting. It is in successful operation in a number of large works, and is being installed in others. Our furnaces have outgrown the shovel; we have passed the limit of efficiency of the old methods of handling material for them. We must come to mechanical feeding in spite of ourselves. But whatever may be the motive leading to its introduction, its chief justification will be discovered, after it has been successfully installed and correctly adjusted, in the consequent great improvement of general operating results, metal saving, etc. It will remove one of the most uncertain factors with which the metallurgist has to deal, thereby bringing into clearer view for study and regulation the other factors (fuel and blast proportion, slag composition, etc.) in a way that has hardly been possible under the irregularities consequent upon hand feeding.
MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES[12]
By Arthur S. Dwight
(January 17, 1903)
Historical.—A silver-lead furnace fed by means of cup and cone was in operation in 1888 at the works of the St. Louis Smelting and Refining Company at St. Louis, Mo., but it is probable that previous attempts had been made, since Hahn refers (“Mineral Resources of the United States,” 1883) in a general way to experiments with this device, which were unsuccessful because the heat crept up in the furnace and gave over-fire. At the time of my visit to the St. Louis works (in 1888) the furnaces were showing signs of over-fire, but this may not have been their characteristic condition. A. F. Schneider, who built the St. Louis furnaces, afterward erected, at the Guggenheim works at Perth Amboy, N. J. , round furnaces with cup and cone feeders, but although good results are said to have been obtained, the running of refinery products is no criterion of what they would do on general ore smelting.
Cup and Cone Feeders.—The cup and cone is an entirely rational device for feeding a round furnace, but is quite unsuitable for feeding a rectangular one. Furnaces of the latter type were installed for copper smelting at Aguas Calientes, Mex., with two sets of circular cup and cone feeders, but disastrous results followed the application of this device to lead furnaces. The reason is clear when it is considered that a circular distribution cannot possibly conform to the requirements of a rectangular furnace. A more rational device was designed for the works at Perth Amboy, N. J.
Fig. 2.—Perth Amboy, N. J. , Lead Furnace. Vertical section at right angles to Fig. 3.
Pfort Curtain.—About ten years ago some of the American smelters adopted the Pfort curtain, which, as adapted to their requirements, consisted of a thimble of sheet iron hung from the iron deck plates so as to leave about 15 in. of space between it and the furnace walls, this space being connected with the down-take of the furnace. The thimble was kept full of ore up to the charge-floor. This device was popular for a time, chiefly because it prevented the furnace from smoking and diminished the labor of feeding, but it was found to give bad results in the furnaces, it being impossible to observe how the charge sunk (except by dropping it below the thimble), while the curtain had to be removed in order to bar down accretions, and, most important, it caused irregular furnace work and high metal losses, because it effected a distribution of the coarse and fine material which was the reverse of correct, the evil being emphasized by the taking off of the gases close to the furnace walls.
Fig. 3.—Perth Amboy, N.J., Lead Furnace. Vertical section at right angles to Fig. 2.
Terhune Gratings.—R. H. Terhune designed a device (United States patent No. 585,297, June 29, 1897), which comprised two grizzlies, one on each side of the furnace, sloping downward from the edge of the charge-floor toward the center line of the furnace. The bars tapered toward the center of the furnace, the open spaces tapering correspondingly toward the sides, so that as the charge was dumped on them a classification of coarse and fine would be effected. This device is correct in conception.
Pueblo System.—In the remodeling of the plant of the Pueblo Smelting and Refining Company in 1895, under the direction of W. W. Allen, mechanical feeding was introduced, and the system was the first one to be applied successfully on a large scale. The furnaces of this plant are 60 × 120 in. at the tuyeres, with six tuyeres, 4 in. in diameter on each side, the nozzles (water cooled) projecting 6 in. inside the jackets. The hight of the smelting column above the tuyeres is 20 ft. The gases are taken off below the charge-floor, and the furnace tops are closed by hinged and counter-weighted doors of heavy sheet iron, opened by the attendant, just previous to dumping the charge-car. In the side walls of the shaft are iron door-frames, ordinarily bricked up, but giving access to the shaft for repairs or barring out without interfering with the movement of the charge-car. Extending across the shaft, about 18 in. above the normal stock line, are three A-shaped cast-iron deflectors, dividing the area of the shaft into four equal rectangles.
The general arrangement of the plant is shown in Fig. 4. From the charge-car pit there extends an inclined trestle, on an angle of 17 deg. to the charge-floor level, in line with the battery of furnaces. The gage of the track is approximately equal to the length of the furnaces at the top. The charge-car, actuated by a steel tail-rope, moves sideways on this track from the charging-pit to any furnace in the battery. The hoisting drums are located at the crest of the incline, inside of the furnace building. At the far end of the latter there is a tightener sheave, with a weight to keep proper tension on the tail-rope. The charge-car has a capacity of 5 tons. It has an A-shape bottom, and is so arranged that one attendant can quickly trip the bolt and discharge the car.
Fig. 4.—Pueblo System. Longitudinal vertical section through incline.
While the car is making its trip the charge-wheelers are filling their buggies, working in pairs, each man weighing up a halfcharge of a particular ingredient. They then separate, each taking his proper place in the line of wheelers on either side. When the car has returned, the wheelers successively discharge their buggies into opposite ends of the car. The coke is added last, to avoid crushing. The system is not strictly economical of labor, since the wheelers, who must always be ready for their car, have to wait for its return, which necessitates more wheelers than would otherwise be required. Figs. 5, 6 and 7 show the car.
Fig. 5.—Pueblo Charge-car. (Side elevation.)
A vertical section through the car filled by dumping from the two ends will show an arrangement of coarse and fine, which is far from regular. Analyzing its structure, we shall find a conical pile near each end, with a valley between them, in which coarse ore will predominate. The deflectors in the furnace, previously referred to, serve to scatter the fines as the charge is dropped in. Without them the feeding of the furnace would be a failure; with them it is successful, though not so completely as might be, the furnaces having a tendency to run with hot tops. With the battery of seven furnaces, each smelting an average of 100 tons of ore per day, the saving, as compared with hand-feeding, was $63 per day, or 9c. per ton of ore, this including cost of steam, but not wear and tear on the machinery. This is distinctly a maximum figure; with fewer furnaces the fixed charges of the mechanical feed would soon increase the cost per ton to such a figure that the two systems would be about equal in economy.
Fig. 6.—Pueblo Charge-car. (Plan.)
Fig. 7.—Pueblo Charge-car. (End elevation.)
East Helena System.—This was introduced at the East Helena plant of the United Smelting and Refining Company by H. W. Hixon. The plant comprised four lead furnaces, each 48 × 136 in., with a 21 ft. smelting column. They were all open-top furnaces, fed through a slot over the center, the gases being taken off below the floor. They were capable of smelting about 180 tons of charge (ore and flux) per 24 hours, using a blast of 30 to 48 oz., furnished by two Allis duplex, horizontal, piston blowers, air-cylinders 36 in. diam., 42 in. stroke, belted from electric motors. The Hixon feed was designed to meet existing conditions, without irrevocably cutting off convenient return to hand feeding in case of an emergency. As shown in Fig. 9 there is a track-way at right angles to the line of furnaces. The car hoisted up the incline is landed on a transfer carriage, on which, after detaching the cable, it can be moved over the tops of the furnaces by means of a tail-rope system. The gage of the charge-car is 4 ft. 9 in.; of the transfer carriage, 11 ft. 8 in. A switch at the lower end of the incline permits two charge-cars to be employed, one being filled while the other is making the trip. In sending down the empty car a hand winch is necessary to start it from the transfer carriage. Figs. 10 and 11 show the charge-car; Fig. 12 the transfer carriage.
Fig. 8.—Pueblo System. (Sectional diagrams of furnace top.)
The charge-car is 10 × 4 × 3.5 ft., and has capacity for 6 tons of ore, flux, slag and fuel, the total of ore and flux being usually 8800 lb. Its bottom is flat, consisting of two doors, hinged along the sides and kept closed by means of chains wound about a longitudinal windlass on top of the car. The charging pits are decked with iron plates, leaving a slot along the center of each car exactly like the slot in the furnace top. The loaded ore-buggies are taken from the wheelers by two men, who carefully distribute the contents of each buggy along the whole length of the charge-car by dragging it along the slot while in the act of dumping. Each buggy contains but one ingredient; they follow one another in a prescribed order, so as to secure thin layers in the charge-car. The coke is divided into three or more layers.
Fig. 9.—East Helena System. (Vert-longitudinal section and plan of incline.)
Fig. 10.—East Helena Charge-car. (Side elevation.)
The first few trials of this device were not satisfactory. The furnaces quickly showed over-fire, and decreased lead output, which would not yield to any remedy except a return to hand feeding. The total charge being dropped in the center of the furnace, a central core of fines was produced, the lumps tending to roll toward the walls. This wrong tendency was emphasized by the presence of the chains supporting the bottom of the charge-car. On unwinding them to dump the car, the doors were prevented from dropping by the wedging of the chains in the charge, which in turn arched itself more or less against the sides of the car; hence the doors opened but slowly, and often had to be assisted by an attendant with a bar. In consequence of this slow opening, considerable fine ore sifted out first and formed a ridge in the center of the furnace, from the slopes of which the coarser part of the charge, the last to fall, naturally rolled toward the sides. This fact, determined during a visit of the writer in April, 1899, proved to be the key to the situation. The attendant operating the tail-rope mechanism was instructed to move the transfer carriage rapidly backward and forward over the slot while the first one-third or one-half of the charge was dropping, and during the rest of the discharge to let the car stand directly over the slot and permit the coarser material to fall in the center of the furnace. Two piles of comparatively fine material were thus left on the charge-floor, one on each side of the slot. These were subsequently fed in by hand, with instructions to throw the material well to the sides of the furnace.
Fig. 11.—East Helena Charge-car. (Plan.)
The furnaces were running very hot on top when this modified procedure was begun. In a few hours the over-fire had disappeared; the lead output was increasing; and the furnaces were running normally. This was done about May 1, 1899, and from that time until about February 20, 1900, the Hixon feed, as modified above, was continuously in operation. In October, 1898, with three furnaces in operation and hand feeding, the labor cost per furnace was $42.06 per day; in October, 1899, with the same number of furnaces and mechanical feeding, it was $41 per day, the saving being only 0.6c. per ton of charge.
Fig. 12.—East Helena Charge-car and Transfer Carriage. (Elevation.)
Fig. 13.—East Helena System, with spreader and curtains. (Experimental form.)
Dwight Spreader and Curtain.—In January, 1900, the writer again had occasion to visit the East Helena plant, to investigate why a certain cheap local coke could not be used successfully instead of expensive Eastern coke. Strange as it may seem, the peculiar behavior of the cokes was traced to improper feeding of the furnaces. Further study of the mechanical feeding system, then in operation for nine months, showed that it was far from perfect, and it appeared desirable to design a spreader which would properly distribute the material discharged from the Hixon car and dispense with hand feeding entirely. An experimental construction was arranged, as shown in Fig. 13. The flanged cast-iron plates around the feeding slot were pushed back and a roof-shaped spreader, with slopes of 45 deg., was set in the gap, leaving openings about 8 in. wide on each side. The plan provided for two iron curtains to be hung, one on each side of the spreader, and so adjusted that the fine ore sliding down the spreader would clear the edge of the curtain and shoot toward the sides of the furnace, while the coarse ore would strike the curtain and rebound toward the center of the furnace. The classification effected in this manner was capable of adjustment by raising or lowering the curtain. This arrangement was found to work surprisingly well. The first furnace equipped with it immediately showed improvement. It averaged better in speed, with lower blast, lower lead in slag and matte, and better bullion output than the other furnaces operating under the old system. The success of the spreader and curtain being established, the furnaces were provided with permanent constructions, the only modifications being that the ridge of the spreader was lowered to correspond with the level of the floor and the curtains were omitted, the feeding being apparently satisfactory without their aid. In their absence, the lowering of the spreader was a proper step, as it distributed the material fully as well, and caused less abrasion of the walls. The final form is shown approximately in Fig. 14. It has given complete satisfaction at East Helena since February, 1900, and has been adopted as the basis for the mechanical feeding device in the new plant of the American Smelting and Refining Company at Salt Lake, Utah.
Fig. 14.—East Helena System. (Final form, approximate.)
Comparison of Systems.—In mechanical design the Pueblo system is better than the East Helena, being simpler in construction and operation. No time is lost in attaching and changing cables, operating transfer carriage, etc. In both systems the track runs directly over the tops of the furnaces, and this is an inconvenience when furnace repairs are under way. The Pueblo car is the simpler, and makes the round trip in about half the time of a car at East Helena, so the two cars of the latter do not make much difference in this respect. The system of filling the charge-car at Pueblo is also the quicker. It may be estimated roughly that per ton of capacity it takes 2.5 to 3 times as long to fill the East Helena car; and this means longer waiting on the part of the wheelers, and consequently greater cost of moving the material, representing probably 7 or 8c., in favor of Pueblo, per ton of charge handled. However, both systems are wasteful of labor. As to furnace results, it is believed that the better distribution of the charge in the East Helena system leads to greatly increased regularity of furnace running, less tendency to over-fire, some economy in fuel, less accretions on the furnace walls and larger metal savings. If the half of these conclusions are true, the difference of 7 or 8c. per ton in favor of the Pueblo system, which can be traced almost entirely to the cost of filling the charge-car, sinks into insignificance in comparison with the important advantages of having the furnaces uniformly and correctly fed.
True Function of the Charge-Car.—The radically essential feature of a mechanical feeding device is that part which automatically distributes the material in the furnace, whatever approximate means may have been used to effect the delivery.
Taking a hasty review of the numerous feeding devices that have been tried in lead-smelting practice, we cannot but remark the fact that those which depended upon dumping the charge into the furnace from small buggies or barrows failed generally to secure a proper classification and distribution of coarse and fine, and, consequently, were abandoned as unsuccessful, while the adoption of the idea of the charge-car for transporting the material to the furnace in large units seems to have been coincident with a successful outcome. It is natural enough, therefore, that the car should be regarded by many as the vital feature. This view of the question is not, however, in accordance with the true perspective of the facts, and merely limits the field of application in an entirely unnecessary way. It must be apparent that the essential function of the charge-car is cheap and convenient transportation. The distribution of the charge is an entirely different matter, in which, however, the charge-car may be made to assist, as in the Pueblo system; or entirely distinct and special means may be employed for the distribution, as in the East Helena system.
To follow the argument to its conclusion, let us imagine for the moment that the East Helena plant were arranged on the terrace system, with the furnace tops on a level with the floor of the ore-bins. Certain precautions being observed, the spreader would give as good results with small units of charge delivered by buggies as it now does with the large units delivered by the charge-car, and the expense of delivery to the furnaces would be practically no more than it now is to the charge-car pit. The furnace top would, of course, have to be arranged so that the buggies, in discharging, could be drawn along the slot, so as to give the necessary longitudinal distribution parallel to the furnace walls, just as is now done in filling the charge-car. The ends of the spreader, if built like a hipped roof, would secure proper feeding of the front and back.
Thus, by eliminating the charge-car, and with it the necessity for powerful hoisting machinery, with its expensive repairs and operating costs, we may greatly simplify the problem of mechanical feeding, and open the way for the adoption of successful automatic feeding in many existing plants where it is now considered impracticable.
COST OF SMELTING AND REFINING
By Malvern W. Iles
(August 18, 1900)
In the technical literature of lead smelting there is a lamentable lack of data on the subject of costs. The majority of writers consider that they have fulfilled their duties if they discuss in full detail the chemical and engineering sides of the subject, leaving the industrial consideration of cost to be wrought out by experience. When an engineer or metallurgist collects data on the costs involved in the various smelting operations, he generally hesitates to give this special information to the public, as he regards it as private, or reserves it as stock in trade to be held for his own use.
The following tables of cost have been compiled from actual results of smelting and refining at the Globe works, Denver, Colo., and are offered in the hope that they will prove a valuable addition to the literature of lead smelting. These results are offered tentatively, and, while true for the periods stated, they require considerable adjustment to meet the smelting conditions of the present time.
COST OF HAND-ROASTING PER TON (2000 LB.) OF ORE
| 1887 | $3.975 | │ | 1893 | — |
| 1888 | 4.280 | │ | 1894 | 3.429 |
| 1889 | 4.120 | │ | 1895 | 2.806 |
| 1890 | 3.531 | │ | 1896 | 2.840 |
| 1891 | 3.530 | │ | 1897 | 2.740 |
| 1892 | — | │ | 1898 | 2.620 |
At first the roasting was done mainly by hand roasters; later two Brown-O’Harra mechanical furnaces were used, and the cost was reduced, but not to the extent usually conceded to this type of furnace, as the large amount of repairs and the consequent loss of time diminished the apparent gain due to greater output. The figures quoted above may be considered somewhat higher than the average, as the roasters were charged in proportion with expenses of general management, office, etc.
In viewing the yearly reduction of costs one must take into consideration many changes in the furnace construction and working, as well as the items of labor, fuel, etc. From 1887 to 1899 the principal changes in the construction of the hand-roasting furnaces consisted in an increase of width, 2 ft., which allowed an addition of 200 lb. to each ore charge, and corresponded to a total increase per furnace of 1200 lb. in 24 hours. In the working of the charge an important change was made in the condition of the product. Formerly the material was fused in the fusion-box and drawn from the furnace in a fused or slagged condition; and while this gave an excellent material for the subsequent treatment in the shaft furnace in that there was very little dusting of the charge, and a considerable increase in the output of the furnace, the disadvantages of large losses of lead and silver greatly over-balanced the advantages, and called for an entire abandonment of the fusion-box. As a result of experience it was found that the best condition of product is a semi-fused or sintered state, in which the particles of roasted ore have been compressed by pounding the material, which has been drawn into the slag pots, with a heavy iron disk. The amount of “fines” under these conditions is quite small and depends upon the percentage of lead in the ore, the degree of heat employed, and the extent of the compression.
The total cost was partly reduced from the lessened labor cost following the financial disturbance of 1893, and partly from the reduction in the fuel cost, the former expensive lump coal being replaced by the slack coals from southern Colorado.
The comparison of the cost of labor by the two methods shows a gain of 54c. a ton in favor of the mechanical furnaces. However, I consider that this gain is a costly one, and is more than offset by the large amount of high-grade fuel required, and the expense of repairs not shown in the following table. Indeed, I believe that at the end of five or ten years the average cost of roasting per ton by the hand roasters will be even smaller than by these mechanical roasters.
To illustrate the details of roasting cost and to furnish a comparison of the hand roasters and mechanical furnaces, the following table has been prepared:
DETAILS OF AVERAGE MONTHLY COST FOR 1898 OF HAND ROASTERS AND MECHANICAL FURNACES
| Month | Hand Roasters | Brown-O’Harra Mechanical Furnaces | ||||||
|---|---|---|---|---|---|---|---|---|
| Total Tons Roasted Per Day | Tons Roasted Per Day | Labor | Coal | General Expense | Labor | Coal | General Expense | |
| January | 5,691 | 184 | $1.47 | $0.53 | $0.80 | $0.92 | $0.80 | $1.32 |
| February | 5,677 | 203 | 1.44 | 0.44 | 0.99 | 0.72 | 0.58 | 1.01 |
| March | 5,821 | 188 | 1.51 | 0.53 | 0.64 | 0.76 | 0.64 | 0.62 |
| April | 5,472 | 182 | 1.47 | 0.47 | 0.71 | 0.80 | 0.69 | 0.87 |
| May | 5,444 | 176 | 1.55 | 0.51 | 0.84 | 0.80 | 0.69 | 0.81 |
| June | 4,859 | 162 | 1.58 | 0.48 | 0.71 | 0.90 | 0.68 | 1.17 |
| July | 5,691 | 184 | 1.59 | 0.48 | 0.75 | 0.72 | 0.56 | 0.64 |
| August | 5,910 | 191 | 1.55 | 0.46 | 0.83 | 0.72 | 0.55 | 0.75 |
| September | 5,677 | 189 | 1.55 | 0.45 | 0.74 | 0.73 | 0.55 | 0.67 |
| October | 6,254 | 202 | 1.48 | 0.49 | 0.72 | 0.65 | 0.50 | 0.60 |
| November | 6,291 | 213 | 1.42 | 0.47 | 0.80 | 0.66 | 0.53 | 0.70 |
| December | 5,874 | 198 | 1.45 | 0.48 | 0.78 | 0.79 | 0.63 | 0.81 |
| Average | $1.50 | $0.48 | $0.77 | $0.76 | $0.62 | $0.83 | ||
| Total | 2.75 | 2.21 | ||||||
Cost of Smelting.—The lead-ore mixtures of the United States, in addition to lead, contain gold, silver and generally copper, and are treated to save these metals. The total cost of smelting is made up of a large number of items. The questions of locality and transportation, fuel, fluxes and labor are the principal factors, to which must be added the handling of the material to and from the furnace; the furnace itself, its size, shape, and method of smelting, the volume and pressure of blast, etc. The following table of costs, from 1887 to 1898, shows in a general way the great advance that has been made in the development of smelting, and the consequent reduction in cost per ton of ore treated:
AVERAGE COST OF SMELTING, PER TON
| 1887 | $4.644 | 1891 | 4.170 | 1895 | 2.786 |
| 1888 | 4.530 | 1892 | 4.906 | 1896 | 2.750 |
| 1889 | 4.480 | 1893 | 3.375 | 1897 | 2.520 |
| 1890 | 4.374 | 1894 | 3.029 | 1898 | 2.260 |
In connection with this table of smelting cost should be considered the changes developed during the interval 1887-1889, outlined as follows:
CONDITIONS OF SMELTING IN 1886 AND 1899 CONTRASTED TO SHOW THE PROGRESS OF DEVELOPMENT
| Area of Furnace at Tuyeres, In. | Height of Charge from Tuyeres, Ft. | Blast Pressure, Lb. per Sq. In. | Fore Hearth Capacity, Cu. Ft. | Slag Settled | Fuel | Slag Removed, Lb. per Trip | Matte Removed, Lb. per Trip | |
|---|---|---|---|---|---|---|---|---|
| 1886 | 30 × 100 | 11 | 1 | 6 | In pots | Charcoal | By hand 280 | By hand 200 |
| 1899 | 42 × 140 | 16 | 3 to 4 | 128 | In furnaces | Coke | By locomotive 3000-6000 | By horse 2000-3000 |
I believe that there is room for further improvement in the substitution of mechanical transportation within the works for hand labor, and that the fuel cost can be materially reduced by replacing the coke, which at present contains 16 to 22 per cent. of ash, by a fuel of purer and better quality.
Cost of Refining by the Parkes Process.—In general it may be stated that the average cost of refining base bullion is from $3 to $5 a ton. This amount is based on the cost of labor, spelter, coal, coke, supplies, repairs and general expenses. When the additional items of interest, expressage, brokerage and treatment of by-products are considered, which go to make up the total refining cost, the amount may be stated approximately as $10 per ton of bullion treated.
Variations in the cost occur from time to time, and are due to several causes, principally the irregularity of the bullion supply and its consequent effect on the work of the plant. When the amount of bullion available for treatment is small, the plant cannot be run to its maximum capacity, and the cost per ton will naturally be increased. To illustrate this variation, the average cost per ton of base bullion refined during nine months in 1893 was:
January, $4.864; February, $5.789; March, $5.024; April, $3.915; May, $5.094; June, $4.168; July, $4.231; August, $4.216; September, $5.299.
The yearly variation shows but little change, as the average cost per ton was for 1893, $4.75; for 1894, $3.99; for 1895, $4.21; for 1896, $3.90. In considering the total cost of refining, the additional factors of interest, expressage, parting, brokerage, and reworking of by-products must be considered. As the doré silver is treated at the works or elsewhere, so will the total cost be less or greater. The following table gives the cost in detail, when the parting is done at the same works:
AVERAGE MONTHLY COST OF REFINING PER TON OF BULLION TREATED
| Items | 1895Jan. to July | 1895July to Dec. | 1896Jan. to July | Average |
|---|---|---|---|---|
| Labor | $2.351 | $1.718 | $1.836 | $1.968 |
| Spelter | 0.757 | 0.840 | 0.987 | 0.861 |
| Coal | 0.585 | 0.442 | 0.461 | 0.496 |
| Coke | 0.634 | 0.418 | 0.511 | 0.521 |
| Supplies, repairs and general expenses | 0.343 | 0.273 | 0.252 | 0.289 |
| Interest | 1.808 | 1.075 | 1.070 | 1.317 |
| Expressage | 1.360 | 1.015 | 0.882 | 1.085 |
| Parting and brokerage | 2.483 | 2.084 | 1.796 | 2.121 |
| Reworking by-products | 1.567 | 1.286 | 1.625 | 1.492 |
| Totals | $11.888 | $9.151 | $9.420 | $10.151 |
| Tons bullion refined 5,511.58 | 9,249.07 | 10,103.43 | 8,287.99 |
An analysis of the different items of cost is important, and a brief summary is given below.
Labor and Attendance.—The cost for this item varies but little from year to year, and its reduction depends, for the most part, on a larger yield per man rather than on a reduction of wages. If a man at the same or slightly increased cost can give a larger output, so will the labor cost per ton be diminished. This result is accomplished by enlarging the furnace capacity and by using appliances which will handle the bullion and its products in an easier and quicker manner. The small size of the furnaces, settlers and retorts used at modern refineries is open to criticism; I believe that great improvement can be made in this direction.
Spelter.—The cost of this item varies with the market conditions, and will probably be changed but little in the future, as the amount necessary per ton of bullion seems to be fixed.
Coal.—The amount required per ton of bullion is fairly constant, and while lessened cost for fuel may be attained by the substitution of oil or gaseous fuel, the fuel cost in comparison with the aggregate cost is very small, and leaves little opportunity for improvement in this line.
Supplies.—This item includes brooms, shovels, wheelbarrows, etc., and the amount is small and fairly constant from year to year.
Repairs.—This item is quite small in works properly constructed; and in this connection I wish to call particular attention to the floor covering, which should be made of cast-iron plates from 1.5 to 2 in. thick, and placed on a 2 to 3 in. layer of sand spread over the well-tamped and leveled ground. The constant patching of brick floors is not only an annoyance, but is costly from the additional labor required. Furthermore, a brick floor does not permit a close saving of the metallic scrap material.
It will be found economical in the long run to protect all exposed brickwork of furnaces or kettles with sheet iron.
In the construction of the refinery building I should advise brick walls except at the end or side, where there is the greatest likelihood of future extension; here corrugated iron may be used. The roof should not be made of corrugated iron, as condensed or leakage water is liable to collect and drop on those places where water should be scrupulously avoided. The presence of water in a mold at the time of casting, even though small in amount, will cause explosions and will scatter the molten lead, endangering the workmen.
The item of repair for the ordinary corrugated iron roof may be diminished by constructing it of 1 in. boards with intervening spaces of half an inch, the whole overlaid with tarred felt, and covered with sheets of iron at least No. 27 B. W. G., painted with graphite paint and joined together with parallel rows of ribbed crimped iron.
General Expenses.—This item is generally constant, and calls for no special comment.
Interest.—This important item is, as a rule, considerable, as the stock of bullion and other gold-and silver-bearing material is quite large. For this reason special attention should be given to prevent the accumulation of stock or by-products. The occasional necessity of additional capital to run the business should preferably be met by an increase of working capital, rather than by a direct loan.
Expressage.—This item, as a rule, is large, and should be taken into consideration in the original plans for the location of the refining works.
Parting.—The item of parting and brokerage is the largest of the refinery costs, and for obvious reasons a modern smelting plant should have a parting plant under its own control.
The Working of the By-Products.—This constitutes a large item of cost, and considerable attention should be devoted to the improvement of present methods, which seem faulty, slow and expensive.
Summary.—The items of smaller cost with their respective amounts per ton of base bullion treated are: Spelter, $0.85; coal, $0.50; coke, $0.50; supplies, repairs and general expenses, $0.35; total, $2.10. It is doubtful whether much improvement can be made in the reduction of these costs.
The items of larger cost are: Labor, $2; interest, $1.32; expressage, $1.10; parting and brokerage, $2; reworking by-products, $1.50; total, $7.92. The general manager usually attends to the items of interest, expressage and brokerage, leaving the questions of labor and working of by-products to the metallurgist.
The cost quoted for smelting practice, as employed at Denver, will differ necessarily from those at other localities, where the cost of labor, freight rates on spelter, fuel, etc., are changed. Refining can doubtless be done at a lower cost at points along the Mississippi River, and even more so at cities on the Atlantic seaboard, as Newark or Perth Amboy, N. J.
The consolidation of many of the more important smelting plants of the United States under one management will doubtless alter the figures of cost given above, particularly as the interest cost there stated is at the high rate of 10 per cent., a condition of affairs now changed to 5 per cent. Other factors have lessened the cost of refining; the bullion produced at the present time is softer, or contains a smaller amount of impurities, and admits of easier working with shorter time and less labor. By proper management larger tonnages are turned out per man, and the Howard stirrer and Howard press have simplified and cheapened the working of the zinc skimmings. To illustrate the comparatively recent conditions of cost I have compiled the following table for each month of the year 1898:
COST OF REFINING DURING 1898, INCLUDING LABOR, SPELTER, COAL, COKE, SUPPLIES, REPAIRS AND GENERAL EXPENSES.
| January | $3.59 | May | 3.38 | September | 3.35 |
| February | 3.28 | June | 3.56 | October | 3.45 |
| March | 3.26 | July | 3.65 | November | 3.20 |
| April | 3.59 | August | 3.54 | December | 3.56 |
| Average cost during the year, $3.45. | |||||
It is understood, of course, that these figures do not include cost of interest, expressage, parting, brokerage and reworking of by-products.
[Although this article refers to conditions in 1898, since which time there have been improvements in practice, the latter have not been of radical character and the figures given are fairly representative of present conditions.—Editor.]
SMELTING ZINC RETORT RESIDUES[13]
By E. M. Johnson
(March 22, 1906)
The following notes were taken from work done at the Cherokee Lanyon Smelter Company, Gas, Kansas, in 1903. It was practically an experiment. The furnace was only 36 × 90 in. at the crucible, with a 10 in. side bosh and a 6 in. end bosh. There were five tuyeres on each side with a 3 in. opening. The side jackets measured 4.5 ft. × 18 in. The distance from top of crucible to center of tuyeres was 11.5 in.
The blast was furnished by one No. 4½ Connellsville blower. The furnace originally was only 11 ft. from the center of tuyeres to the feed-floor, and had only been saving about 60 per cent. of the lead. This loss of lead, however, was not entirely due to the low furnace. As no provision had been made to separate the slag and matte, upon assuming charge I raised the feed-floor 3 ft., thereby changing the distance from the tuyere to top of furnace from 11 ft. to 14 ft. Matte settlers were also installed. These two changes raised the percentage of lead saved to 92, as shown by monthly statements. The furnace being small, and a high percentage of zinc oxide on the charge, the campaigns were naturally short. The longest run was about six weeks. This was made on some residue that had been screened from the coarse coal, and coke, and had weathered for several months. This particular residue also carried about 10 per cent. lead. The more recent residue that had not been screened and weathered, and was low in lead, did not work so well. Although these residues consisted of a large proportion of coal and coke, it seemed impossible to reduce the percentage of good lump coke on the charge lower than 12.5 or 13 per cent. At the same time the reducing power of the residue was strong, and with the normal amount of coke caused some trouble in the crucible.
When residue containing semi-anthracite coal was smelted, the saving in lead dropped, and the fire went to the top of the furnace, burning with a blue flame, thereby necessitating the reduction of this class of material. This residue had been screened through a five-mesh screen, and wet down in layers, becoming so hard that it had to be blasted. The low saving of lead with this class of material was a surprise, as it has been claimed that the substitution of part of the fuel by anthracite coal did not affect the metallurgical operations of the furnace.
The slag was quite liquid and flowed very well at all times. However, there was a marked variation in the amount at different tappings. This, I am satisfied, was not due to irregular work on the furnace, but may be accounted for in the following manner. The residue (not screened or weathered to any extent), consisting approximately of one-half coal and coke, was very bulky, and while there was about 35 per cent. of it on the charge by weight, there was over 50 per cent. of it by bulk, not including slag and coke. In feeding, therefore, it was a difficult matter to mix the whole of it with the charge. Several different ways of feeding the furnace were tried. The one giving the most satisfactory results was to feed nearly all of the residue along the center of the furnace, in connection with the lime-rock, coarse ore and coarse iron ore, and the fine and easy smelting ores along the sides. The slag was spread uniformly over the whole furnace, while the sides were favored with the coke. The charge would drop several inches at a time, going down a little faster in the center than on the sides.
It is possible that a small proportion of the residue in connection with the easy smelting, leady, neutral ore, iron ore and lime-rock formed the type of slag marked No. 1.
| SiO2 | FeO | MnO | CaO | ZnO | Pb | Ag | |
|---|---|---|---|---|---|---|---|
| 1 | 33.7 | 34.1 | 1.0 | 16.5 | 7.5 | 0.9 | 0.7 |
| 2 | 31.0 | 36.1 | 1.2 | 16.0 | 9.6 | 1.3 | — |
This being tapped with a good flow of slag, the charge would drop, bringing a proportionately large amount of residue in the fusion zone which formed the type of slag marked No. 2. There was also a marked variation in the slag-shells from different pots. The above cited irregularities of course exist to a certain extent in any blast furnace.
AVERAGE ANALYSIS OF MATERIALS SMELTED
| Name | SiO2 | FeO | CaO | MgO | ZnO | Al2O3 | Fe2O3 | S | Pb | Cu | Ag | Au |
|---|---|---|---|---|---|---|---|---|---|---|---|---|
| Mo. iron ore | 10.0 | 65.0 | ; | ; | ; | ; | ; | ; | ; | ; | ; | ; |
| Lime rock | 1.5 | ; | 52.0 | ; | ; | ; | ; | ; | ; | ; | ; | ; |
| Mo. galena | 1.5 | 2.4 | ; | ; | 9.5 | ; | ; | 11.0 | 74.0 | ; | ; | ; |
| Av. of beds | 50.8 | 16.2 | ; | ; | 4.6 | ; | ; | 3.3 | 9.1 | ; | ; | ; |
| Residue[14] | 10.5 | 38.5 | ; | ; | 18.0 | ; | ; | 4.8 | 2.2 | 1.0 | 10.0 | 0.03 |
| Roasted matte[15] | 9.0 | 48.0 | 3.0 | ; | 10.0 | ; | ; | 4.0 | 9.9 | 3.0 | 21.0 | 0.06 |
| Barrings | 18.8 | 24.4 | 5.0 | ; | 14.5 | ; | ; | 6.0 | 25.4 | ; | 13.0 | 0.07 |
| Coke ash | 27.0 | ; | 14.9 | 4.5 | ; | 19.7 | 31.6 | ; | ; | ; | ; | ; |
| H2O | V.M. | F.C. | Ash | S | ; | ; | ; | ; | ; | ; | ; | |
| Coke[16] | 1.2 | 2.3 | 85.7 | 11.1 | 0.9 | ; | ; | ; | ; | ; | ; | ; |
ANALYSIS OF BULLION, SLAG AND MATTE PRODUCED
| Bullion | Slag | Matte | |||||||||||
|---|---|---|---|---|---|---|---|---|---|---|---|---|---|
| Ag | Au | SiO2 | FeO | MnO | CaO | ZnO | Pb | Ag | Ag | Au | Pb | Cu | |
| Feb. | 90.0 | 1.15 | 31.2 | 35.9 | 1.0 | 14.5 | 10.3 | 0.88 | 0.98 | 19.0 | 0.04 | 8.7 | 1.5 |
| March | 93.1 | 1.63 | 31.3 | 37.2 | 1.0 | 13.9 | 11.1 | 0.71 | 1.30 | 21.0 | 0.06 | 8.0 | 2.5 |
| April | 104.3 | 1.59 | 29.8 | 37.7 | 2.7 | 13.9 | 11.4 | 0.52 | 1.40 | 23.0 | 0.07 | 7.0 | 3.5 |
| May | 90.0 | 1.24 | 30.0 | 37.3 | 2.2 | 14.1 | 9.3 | 0.86 | 1.10 | 25.4 | 0.07 | 5.1 | 4.0 |
| July | 78.7 | 1.00 | 32.2 | 37.4 | 1.0 | 13.9 | 9.8 | 0.50 | 1.15 | 21.3 | 0.03 | 8.9 | 4.0 |
| Aug. | 90.8 | 1.21 | 31.2 | 37.1 | 1.7 | 13.7 | 9.6 | 1.10 | 1.60 | 23.1 | 0.08 | 9.8 | 3.0 |
| Sept. | 65.3 | 2.58 | 32.0 | 39.7 | 0.8 | 14.1 | 8.1 | 0.80 | 1.30 | 18.6 | 0.06 | 7.6 | 2.3 |
| Average | 87.5 | 1.49 | 31.1 | 37.5 | 1.5 | 14.1 | 10.0 | 0.77 | 1.26 | 21.6 | 0.06 | 7.8 | 3.0 |
MONTHLY RECORD OF FURNACE OPERATIONS
| Blast Ounces | Tons per F. D. | Per Cent. Pb. on Charge | Per Cent. Coke on Charge | Per Cent. Slag on Charge | Per Cent. S on Charge | Matte Produced | Saving | |||
|---|---|---|---|---|---|---|---|---|---|---|
| Ag | Au | Pb | ||||||||
| Feb. | 21 | 42.5 | 9.0 | 12.0 | 30.0 | 3.7 | 8.0} | 84.4 | 83.0 | 90.3 |
| March | 21 | 44.8 | 9.7 | 13.5 | 37.0 | 4.0 | 9.0} | |||
| April | 21 | 43.7 | 9.0 | 13.5 | 35.0 | 4.3 | 10.0 | 97.9 | 70.5 | 96.6 |
| May | 21 | 49.4 | 10.0 | 13.5 | 30.0 | 3.5 | 6.5 | 95.6 | 109.5 | 88.8 |
| July | 17 | 41.0 | 9.8 | 12.5 | 34.0 | 3.8 | 6.0 | 97.9 | 90.0 | 92.9 |
| August | 18 | 47.0 | 9.3 | 13.0 | 32.0 | 3.7 | 6.3 | 86.2 | 107.5 | 87.6 |
| Sept.[17] | 15 | 51.0 | 7.3 | 13.0 | 30.0 | 2.8 | 4.6 | 92.9 | 94.0 | 95.6 |
| Average | 45.6 | 9.1 | 13.0 | 32.6 | 3.7 | 7.2 | 90.8 | 92.4 | 92.0 | |
I believe that, in smelting residues high in zinc oxide, better metallurgical results would be obtained by using a dry silicious ore in connection with a high-grade galena ore, provided the residue be low in sulphur. This was confirmed to a certain degree in actual practice, as the furnace worked very well upon increasing the percentage of Cripple Creek ore on the charge. This would also seem to indicate that alumina had no bad effect on a zinky slag.
ZINC OXIDE IN SLAGS
By W. Maynard Hutchings
(December 24, 1903)
From time to time, in various articles and letters on metallurgical subjects in the Engineering and Mining Journal, the question of the removal of zinc oxide in slags is referred to, and the question is raised as to the form in which it is contained in the slags.
I gather that opinion is divided as to whether zinc oxide enters into the slags as a combined silicate, or whether it is simply carried into them in a state of mechanical mixture.
For many years I have taken great interest in the composition of slags, and have studied them microscopically and chemically. The conclusion to which I have been led as regards zinc oxide is, that in a not too basic slag it is originally mainly, if not wholly, taken up as silicate along with the other bases. On one occasion, one of my furnaces for several days produced a slag in which beautiful crystals of willemite were very abundant, both free in cavities and also imbedded throughout the mass of solid slag, as shown in thin sections under the microscope. In the same slag was a large amount of magnetite, all of which contained a considerable proportion of zinc oxide combined with it. Magnetite crystals, separated out from the slag and treated with strong acid, yielded shells of material retaining the form of the original mineral, rich in zinc oxide; an inter-crystallized zinc-iron spinel, in fact. I have seen and separated zinc-iron spinels very rich in zinc oxide from other slags. They have been seen in the slags at Freiberg; and of course everybody knows the very interesting paper by Stelzner and Schulze, in which they described the beautiful formations of spinels and willemite in the walls of the retorts of zinc works.
I think there is thus good ground for concluding that zinc oxide is slagged off as combined silicate, and that free oxide does not exist in slags; though zinc oxide does occur in them after solidification, combined with other oxides, in forms ranging from a zinkiferous magnetite to a more or less impure zinc-iron, or zinc-iron-alumina spinel, these minerals having crystallized out in the earlier stages of cooling.
The microscope showed that the crystals of willemite, mentioned above, were the first things to crystallize out from the molten slag. The main constituent was well-crystallized iron-olivine-fayalite.