OTHER METHODS OF SMELTING
THE BORMETTES METHOD OF LEAD AND COPPER SMELTING[43]
BY ALFREDO LOTTI
(September 30, 1905)
It is well known that, in order to obtain a proper fusion in lead
and copper ore-smelting, it is not only advantageous, but often
indispensable, that a suitable proportion of slag be added to the
charge. In the treatment of copper matte in the converter, the total
quantity of slag must be resmelted, inasmuch as it always retains a
notable quantity of the metal; while in the smelting of lead ore in the
blast furnace, the addition of slag is mainly intended to facilitate
the operation, avoiding the use of strong air pressure and thus
diminishing the loss of lead. The proportion of slag required sometimes
amounts to 30 to 35 per cent. of the weight of the ore.
Inasmuch as the slag is usually added in lump form, cold, its original
heat (about 400 calories per kilogram) is completely lost and an
intimate mixture with the charge cannot be obtained. For this reason, I
have studied the agglomeration of lead and copper ores with fused slag,
employing a variable proportion according to the nature of the ore
treated. In the majority of cases, and with some slight modifications
in each particular case, by incorporating the dry or slightly moistened
mineral with the predetermined quantity of liquid slag, and by rapidly
stirring the mixture so as to secure a proper subdivision of the slag
and the mineral, there is produced a spongy material, largely composed
of small pieces, together with a simultaneous evolution of dense fumes
of sulphur, sulphur dioxide, and sulphur trioxide. By submitting this
spongy material to an air blast, the sulphur of the mineral is burned,
the temperature rising in the interior of the mass to a clear red
heat. Copious fumes of sulphur dioxide and trioxide are given off,
and at times a yellowish vapor of sulphur, which condenses in drops,
especially if the ore is pyritous.
At the end of from one to three hours, according to the quantity of
sulphur contained in the material under treatment and the amount of the
air pressure, the desulphurization of the ore, so far as it has come
in contact with the air, is completed, and the mass, now thoroughly
agglomerated, forms a spongy but compact block. It is then only
necessary to break it up and smelt it with the requisite quantity of
flux and coke. The physical condition of the material is conducive to
a rapid and economical smelting, while the mixture of the sulphide,
sulphate and oxide leads to a favorable reaction in the furnace.
In employing this method, it sometimes happens that ores rich in
sulphur produce during the smelting a little more matte than when the
ordinary system of roasting is employed. In such instances, in order
to avoid or to diminish the cost of re-treatment of the matte, it is
best to agglomerate a portion thereof with the crude mineral and the
slag. This has the advantage of oxidizing the matte, which acts as a
ferruginous flux in the smelting.
The system described above leads to considerable economy, especially
in roasting, as the heat of the scoria, together with that given off
in the combustion of the sulphur, is almost always sufficient for the
agglomeration and desulphurization of the mineral; while, moreover, it
reduces the cost of smelting in the blast furnace. Although the primary
desulphurization is only partial (about 50 per cent.), it continues
in the blast furnace, since the mineral, agglomerated with the slag,
assumes a spongy form and thereby presents an increased surface to the
action of the air. The sulphur also acts as a fuel and does not produce
an excessive quantity of matte.
The system will prove especially useful in the treatment of
argentiferous lead ore, since, by avoiding the calcination in a
reverberatory furnace, loss of silver is diminished. It appears,
however, that, contrary to the reactions which occur in the
Huntington-Heberlein process, a calcareous or basic gangue is not
favorable to this process, if the proportion be too great.
The following comparison has been made in the case of an ore containing
62 to 65 per cent. of lead, 16 to 17 per cent. sulphur, 10 to 11 per
cent. zinc, 0.4 per cent. copper, and 0.222 per cent. silver, in which
connection it is to be remarked that, in general, the less zinc there
is in the ore the better are the results.
[Illustration: FIG. 21.—Elevation and Plan of Converting Chambers.]
_Ordinary Method._—Roast-reduction. Cost per 1000 kg. of crude ore:
1. Roasting in reverberatory furnace:
Labor $0.70
Fuel 1.50
Repairs and supplies .05
————- $2.25
2. Smelting in water-jacket:
Labor $1.01
Fuel 2.20
Repairs and supplies .03
Fluxes .50
————- 3.74
————-
Total $5.99
_Bormettes Method._—Agglomeration with slag, pneumatic desulphurization
and smelting in water-jacket:
1. Agglomeration and desulphurization:
Labor $0.42
Repairs and supplies 0.05
———- $0.47
2. Smelting in water-jacket:
Labor $0.90
Fuel 1.91
Repairs and supplies .03
Fluxes .42
————- 3.26
————-
Total $3.73
This shows a difference in favor of the new method of $2.26 per ton of
ore, without taking into account the savings realized by a much more
speedy handling of the operation, which would further reduce the cost
to approximately $2.50 per ton.
[Illustration: FIG. 22.—Details of Transfer Cars.]
In the above figures, no account has been taken of general expenses,
which per ton of ore are reduced because of the greater rapidity of the
process, enabling a larger quantity of ore to be smelted in a given
time. Making allowance for this, the saving will amount to an average
of $2.40 per 1000 kg., a figure which will naturally vary according
to the prices for fuel, labor, and the quantity of matte which it may
be necessary to re-treat. If the quantity of matte does not exceed
10 per cent. of the weight of the ore, it can be desulphurized by
admixture with the ore, without use of other fuel. If, however, the
proportion of matte rises to 20 parts per 100 parts of ore (a maximum
which ought not to be reached in good working), it is necessary to
roast a portion of it. Under unfavorable conditions, consequently,
the saving effected by this process may be reduced to $2 @ $2.20 per
1000 kg., and even to as little as $1.40 @ $1.60. The above reckonings
are, however, without taking any account of the higher extraction of
lead and silver, which is one of the great advantages of the Bormettes
process.
[Illustration: FIG. 23.—Latest Form of Converter. (Section on A B.)]
The technical results obtained in the smelting of an ore of the above
mentioned composition are as follows:
────────────────────────────────────┬─────────────┬─────────────
│ ORDINARY │ BORMETTES
│ METHOD │ METHOD
────────────────────────────────────┼─────────────┼─────────────
Coke, per cent. of the charge │ 14 │ 12
Blast pressure, water gage │12 to 20 cm. │12 to 14 cm.
Tons of charge smelted per 24 hr │ 20 │ 25
Tons of ore smelted per 24 hr │ 8 │ 10
Lead assay of slag │0.80 to 0.90%│0.20 to 0.40%
Matte-fall, per cent. of ore charged│ 5 to 10 │ 10 to 15
Lead extraction │ 90% │ 92%
Silver extraction │ 95% │ 98%
────────────────────────────────────┴─────────────┴─────────────
[Illustration: FIG. 24.—Latest Form of Converter. (Section on C D.)]
The higher extractions of lead and silver are explained by the fact
that the loss of metals in roasting is reduced, while, moreover, the
slags from the blast furnace are poorer than in the ordinary process
of smelting. The economy in coke results from the greater quantity of
sulphur which is utilized as fuel, and from the increased fusibility of
the charge for the blast furnace.
The new system of desulphurization enables the charge to be smelted
with a less quantity of fresh flux, by the employment in its place of a
greater proportion of foul slag. The reduction in the necessary amount
of flux is due not only to the increased fusibility of the agglomerated
charge, but principally to the fact that in this system the formation
of silicates of lead (which are produced abundantly in ordinary
slag-roasting) is almost nil. It is therefore unnecessary to employ
basic fluxes in order to reduce scorified lead.
[Illustration: FIG. 25.—Latest Form of Converter. (Plan.)]
The losses of metal in the desulphurization are less than in the
ordinary method, because the crude mineral remains only a short time
(from one to three hours) in the apparatus for desulphurization and
agglomeration, and the temperature of the process is lower. The
blast-furnace slags are poorer, because there is no formation of
silicate of lead during the agglomeration.
The Bormettes method, in so far as the treatment of lead ore is
concerned, may be considered a combination process of roast-reaction,
of roast-reduction, and of precipitation-smelting. It is not, however,
restricted to the treatment of lead ore. It may also be applied
to the smelting of pyritous copper-bearing ores. In an experiment
with cupriferous pyrites, containing 20 to 25 per cent. sulphur, it
succeeded in agglomerating and smelting them without use of any fuel
for calcination, effecting a perfect smelting, analogous to pyrite
smelting, with the production of a matte of sufficient degree of
concentration.
The first cost of plant installation is very much reduced by the
Bormettes method, inasmuch as the ordinary roasting furnaces are almost
entirely dispensed with, apparatus being substituted for them which
cost only one-third or one-fourth as much as ordinary furnaces. The
process presents the advantage, moreover, of being put into immediate
operation, without any expenditure of excess fuel.
The apparatus required in the process is illustrated in Figs. 21-25.
The apparatus for desulphurization and agglomeration consists of a
cast-iron box, composed of four vertical walls, of which two incline
slightly toward the front. These inclined walls carry the air-boxes.
The other two walls are formed, the one in front by the doors which
give access to the interior, and the other in the rear by a straight
plate. The whole arrangement is surmounted by a hood. The four pieces
when assembled form a box without bottom. Several of these boxes
are combined as a battery. The pots in which the agglomeration and
desulphurization are effected are moved into these boxes on suitable
cars, in the manner shown in the first engraving. A later and more
improved form is shown, however, in Figs. 23-25.
This process, which is the invention of A. Lotti and has been patented
in all the principal countries, is in successful use at the works of
the Société Anonyme des Mines de Bormettes, at Bormettes, La Londe
(Var), France. Negotiations are now in progress with respect to its
introduction elsewhere in Europe.
THE GERMOT PROCESS[44]
BY WALTER RENTON INGALLS
(November 1, 1902)
According to F. Laur, in the _Echo des Mines_ (these notes are
abstracted from _Oest. Zeit._, L., xl, 55, October 4, 1902), A. Germot,
of Clichy, France, made experiments some years ago upon the production
of white lead directly from galena. These led Catelin to attempt the
recovery of metallic lead in a similar way. If air be blown in proper
quantity into a fused mass of lead sulphide the following reaction
takes place:
2PbS + 2O = SO₂ + Pb + PbS.
Thus one-half of the lead is reduced, and it is found collects all the
silver of the ore; the other half is sublimed as lead sulphide, which
is free from silver. The reaction is exothermic to the extent that
the burning of one-half the sulphur of a charge should theoretically
develop sufficient heat to volatilize half of the charge and smelt the
other half. This is almost done in practice with very rich galena,
but not so with poorer ore. The temperature of the furnace must be
maintained at about 1100 deg. C. throughout the whole operation, and
there are the usual losses of heat by radiation, absorption by the
nitrogen of the air, etc. Deficiencies in heat are supplied by burning
some of the ore to white lead, which is mixed with the black fume
(PbS) and by the well-known reactions reduced to metal with evolution
of sulphur dioxide. The final result is therefore the production of
(1) pig lead enriched in silver; (2) pig lead free from silver; (3) a
leady slag; and (4) sulphur dioxide. In the case of ores containing
less than 75 per cent. Pb the gangue forms first a little skin and
then a thick hard crust which soon interferes with the operation,
especially if the ore be zinkiferous. This difficulty is overcome by
increasing the temperature or by fluxing the ore so as to produce a
fusible slag. A leady slag is always easily produced; this is the only
by-product of the process. The theoretical reaction requires 600 cu. m.
of air, assuming a delivery of 50 per cent. from the blower, and at one
atmosphere pressure involves the expenditure of 18 h.p. per 1000 kg. of
galena per hour.
[Illustration: FIG. 26.—Plan and Elevation of Smelting Plant at Clichy.]
The arrangement of the plant at Clichy is shown diagrammatically in
Fig. 26. There is a round shaft furnace, 0.54 meter in diameter and
4.5 meters high. Power is supplied to the blower C through the pulley
G and the shaft DD. The compressed air is accumulated in the reservoir
R, whence it is conducted by the pipe to the tuyere which is suspended
inside of the furnace by means of a chain, whereby it can be raised
or lowered. O₁ and O₂ are tap-holes. L is a door and N an
observation tube. A is the charge tube. X is the pipe which conveys the
gas and fume to the condensation chambers. T is the pipe through which
the waste gases are drawn. V is the exhauster and S is the chimney.
K₁ and K₂ are tilting crucible furnaces for melting lead and
galena.
After the furnace has been properly heated, 100 kg. of lead melted in
K₁ are poured in through the cast-iron pipe P, and after that about
200 kg. of pure, thoroughly melted galena from K₂. Ore containing 70
to 80 per cent. Pb must be used for this purpose. The blast of air is
then introduced into the molten galena, and from 1000 to 3000 kg. of
ore is gradually charged in through the tube A. During this operation
black fume (PbS) collects in the condensation chamber. All outlets are
closed against the external air. If the air blast is properly adjusted,
nothing but black fume is produced; if it begins to become light
colored, charging is discontinued and the blast of air is shut off.
Lead is then tapped through O₂, which is about 0.2 meter above the
hearth, so there is always a bath of lead in the bottom of the furnace;
but it is advisable now and then to tap off some through O₁, so as
gradually to heat up the bottom of the furnace. Hearth accretions are
also removed through O₁. The lead is tapped off through O₂ until
matte appears. The tap hole is then closed, the tuyere is lowered and
the blast is turned into the lead in order to oxidize it and completely
desulphurize the sulphur combinations, which is quickly done. The
oxide of lead is scorified as a very fusible slag, which is tapped off
through O₂, and more ore is then charged in upon the lead bath and
the cycle of operations is begun again.