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.
[Illustration: 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.
[Illustration: 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.
[Illustration: 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.
[Illustration: 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.
[Illustration: FIG. 6.—Pueblo Charge-car. (Plan.)]
[Illustration: 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.
[Illustration: 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.
[Illustration: FIG. 9.—East Helena System. (Vert-longitudinal section
and plan of incline.)]
[Illustration: 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.
[Illustration: 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.
[Illustration: FIG. 12.—East Helena Charge-car and Transfer Carriage.
(Elevation.)]
[Illustration: 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.
[Illustration: 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
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.