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.

_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.

_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.

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.

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.

_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.

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.

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.

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.

_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.

_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 1888 4.280 1889 4.120 1890 3.531 1891 3.530 1892 1893 1894 3.429 1895 2.806 1896 2.840 1897 2.740 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

───────────┬────────┬───────┬─────────────────────┬───────────────────── │ │ │ HAND ROASTERS │ BROWN-O’HARRA │ │ │ │ MECHANICAL FURNACES │ TOTAL │ TONS ├──────┬──────┬───────┼──────┬──────┬─────── Month │ TONS │ROASTED│LABOR │ COAL │GENERAL│LABOR │COAL │GENERAL │ROASTED │PER DAY│ $ │ $ │EXPENSE│ $ │ $ │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 1888 4.530 1889 4.480 1890 4.374 1891 4.170 1892 4.906 1893 3.375 1894 3.029 1895 2.786 1896 2.750 1897 2.520 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│ HEIGHT OF │BLAST PRESSURE,│ FORE HEARTH │ │ AT TUYERES, │CHARGE FROM │ LB. PER │CAPACITY, CU.│ │ IN. │TUYERES, FT.│ SQ. IN. │ FT. │ ────┼───────────────┼────────────┼───────────────┼─────────────┤ 1886│ 30 × 100 │ 11 │ 1 │ 6 │ │ │ │ │ │ │ │ │ │ │ 1899│ 42 × 140 │ 16 │ 3 to 4 │ 128 │ │ │ │ │ │ ────┴───────────────┴────────────┴───────────────┴─────────────┘

────┬────────────┬────────┬───────────────┬───────────────┐ │ SLAG │ FUEL │ SLAG REMOVED, │MATTE REMOVED, │ │ SETTLED │ │ LB. PER TRIP │ LB. PER │ ────┼────────────┼────────┼───────────────┼───────────────┤ 1886│ │ │ │ TRIP │ │ In pots │Charcoal│ By hand │ By hand │ │ │ │ 280 │ 200 │ 1899│ │ │ │ │ │In furnaces │ Coke │ By locomotive │ By horse │ │ │ │ 3000-6000 │ 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 │ 1895 │ 1895 │ 1896 │ AVERAGE │JAN. TO JULY│JULY TO DEC.│JAN. TO JULY│ ─────────────────────┼────────────┼────────────┼────────────┼───────── 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 February 3.28 March 3.26 April 3.59 May 3.38 June 3.56 July 3.65 August 3.54 September 3.35 October 3.45 November 3.20 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 ROW NAME ROW

Mo. iron ore A Roasted matte[15] F Lime rock B Barrings G Mo. galena C Coke ash H Av. of beds D Coke[16] J Residue[14] E

────┬──────┬─────┬────┬────┬────┬─────┬─────┬────┬────┬───┬────┬──── │ SiO₂ │ FeO │CaO │MgO │ZnO │Al₂O₃ │Fe₂O₃ │ S │ Pb │Cu │ Ag │ Au ────┼──────┼─────┼────┼────┼────┼─────┼─────┼────┼────┼───┼────┼──── A │ 10.0 │ 65.0│ │ │ │ │ │ │ │ │ │ B │ 1.5 │ │52.0│ │ │ │ │ │ │ │ │ C │ 1.5 │ 2.4│ │ │ 9.5│ │ │11.0│74.0│ │ │ D │ 50.8 │ 16.2│ │ │ 4.6│ │ │ 3.3│ 9.1│ │ │ E │ 10.5 │ 38.5│ │ │18.0│ │ │ 4.8│ 2.2│1.0│10.0│0.03 F │ 9.0 │ 48.0│ 3.0│ │10.0│ │ │ 4.0│ 9.9│3.0│21.0│0.06 G │ 18.8 │ 24.4│ 5.0│ │14.5│ │ │ 6.0│25.4│ │13.0│0.07 H │ 27.0 │ │14.9│ 4.5│ │ 19.7│ 31.6│ │ │ │ │ │ H₂O │ V.M.│F.C.│ Ash│ S │ │ │ │ │ │ │ J │ 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 ├──────┼────┼─────┼────┼───┼────┼────┼────┼────┼────┼────┼───┼─── Avge.│ 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 │ TONS │PER CENT.│PER CENT.│PER CENT.│PER CENT.│ │OUNCES│ PER │ PB. ON │ COKE ON │ SLAG ON │ S ON │ │ │ F.D. │ CHARGE │ CHARGE │ CHARGE │ CHARGE │ ─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤ Feb. │ 21 │ 42.5 │ 9.0 │ 12.0 │ 30.0 │ 3.7 │ March │ 21 │ 44.8 │ 9.7 │ 13.5 │ 37.0 │ 4.0 │ April │ 21 │ 43.7 │ 9.0 │ 13.5 │ 35.0 │ 4.3 │ May │ 21 │ 49.4 │ 10.0 │ 13.5 │ 30.0 │ 3.5 │ July │ 17 │ 41.0 │ 9.8 │ 12.5 │ 34.0 │ 3.8 │ August │ 18 │ 47.0 │ 9.3 │ 13.0 │ 32.0 │ 3.7 │ Sept.[17]│ 15 │ 51.0 │ 7.3 │ 13.0 │ 30.0 │ 2.8 │ ─────────┼──────┼───────┼─────────┼─────────┼─────────┼─────────┤ Average │ │ 45.6 │ 9.1 │ 13.0 │ 32.6 │ 3.7 │ ─────────┴──────┴───────┴─────────┴─────────┴─────────┴─────────┘

────────┬────────┬─────────────────────┐ │ MATTE │ SAVING │ │PRODUCED│ AG AU PB │ ────────┼────────┼──────┬───────┬──────┤ Feb. │ 8.0} │ 84.4 │ 83.0 │ 90.3 │ March │ 9.0} │ │ │ │ April │ 10.0 │ 97.9 │ 70.5 │ 96.6 │ May │ 6.5 │ 95.6 │ 109.5 │ 88.8 │ July │ 6.0 │ 97.9 │ 90.0 │ 92.9 │ August │ 6.3 │ 86.2 │ 107.5 │ 87.6 │ Sept. │ 4.6 │ 92.9 │ 94.0 │ 95.6 │ ────────┼────────┼──────┼───────┼──────┤ Average │ 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.