│ │ │ LEAD-POISONING │ CASES
│ │ ├─────────────┬────────────────┤ CONTRACTED
│ │ │NO. OF CASES │DAYS OF SICKNESS│AT REVER.│ AT
─────────┼──────┼──────┼─────┬───────┼───────┬────────┤ AND │H. H. SICKNESS
METHOD OF│ YEAR │NO. OF│TOTAL│PER 100│ TOTAL │PER 100 │ SINT. │ FUR.
WORKING │ │ MEN │ │PERSONS│ │PERSONS │ FUR. │
─────────┼──────┼──────┼─────┼───────┼───────┼────────┼─────────┼─────
│ │ │ │ │ │ │ │
Old │{ 1902│ 93 │ 15 │ 16.1 │ 246 │ 264.5 │ 11 │ 4
│{ 1903│ 86 │ 12 │ 13.9 │ 222 │ 258.1 │ 7 │ 5
│ │ │ │ │ │ │ │
H.-H. │ 1904│ 87 │ 8 │ 9.2 │ 242 │ 278.2 │ 6 │ 2
─────────┴──────┴──────┴─────┴───────┴───────┴────────┴─────────┴─────
This shows a gratifying decrease in the number of cases, namely, from
16.1 to 9.2 per cent.; this decrease would have been still greater if
Huntington-Heberlein furnaces had been in use exclusively. However,
most of the time two or three sintering furnaces were fired for
working up by-products, 16 to 18 men being engaged on that work. The
Huntington-Heberlein furnaces alone (at which, in the year 1904, 69 men
in all were occupied) show only 2.9 per cent. of cases. That the number
of days of illness was not reduced is due to the fact that the cases
among the gang of men working at the sintering furnaces were mostly of
long standing and took some time to cure.
The noxious effects upon the health of the workmen in running the
shaft furnaces are due to the fumes from the products made in this
operation, such as work-lead, matte and slag, which flow out of the
furnace at a temperature far above their melting points. Even with
the old method of running the shaft furnaces the endeavor has always
been to provide as efficiently as possible against the danger caused
by this volatilization, and, wherever feasible, to install safety
appliances to prevent the escape of lead vapors into the work-rooms;
but these measures could not be made as thorough as in the case of the
Huntington-Heberlein process.
The principal work in running the shaft furnaces, aside from the
charging, consists in tapping the slag and pouring out the work-lead.
Other unpleasant jobs are the barring down (which in the old process
had to be done frequently) and the cleaning out of the furnace after
blowing out.
In the old process the slag formed in the furnace flows out
continuously through the tap-hole into iron pots placed in front of
the spout. A number of such pots are so arranged on a revolving table
that as soon as one is filled the next empty can be brought up to the
duct; thus the slag first poured in has time to cease fuming and to
solidify before it is removed. The vapors arising from the slag as it
flows out are conveyed away through hoods. At the same time with the
slag, lead matte also issues from the furnace. Now the greater the
quantity of lead matte, the more smoke is also produced; and, with
the comparatively high proportion of lead matte resulting from the
old process, the quantity of smoke was so great that the ventilation
appliances were no longer sufficient to cope with it, thus allowing
vapors to escape into the work-room.
The work-lead collects at the back of the furnace in a well, from which
it is from time to time ladled into molds placed near by. If the lead
is allowed to cool sufficiently in the well, it does not fume much in
the ladling out. But when the furnace runs very hot (which sometimes
happens), the lead also is hotter and is more inclined to volatilize.
In this event the danger of lead-poisoning is very great, for the
workman has to stand near the lead sump.
A still greater danger attends the work of barring down and cleaning
out the furnace. The barring down serves the purpose of loosening
the charge in the zone of fusion; at the same time it removes any
crusts formed on the sides of the furnace, or obstructions stopping
up the tuyeres. With the old furnaces, and their strong tendency to
crust, this work had to be undertaken almost every day, the men being
compelled to work for rather a long time and often very laboriously
with the heavy iron tools in the immediate neighborhood of the glowing
charge, the front of the furnace being torn open for this purpose. In
this operation they were exposed without protection to the metallic
vapors issuing from the furnace, inasmuch as the ventilating appliances
had to be partially removed during this time, in order to render it at
all possible to do the work.
In a similar manner, but only at the time of shutting down a shaft
furnace, the cleaning out (that is to say, the withdrawing of no
longer fused but still red-hot portions of the charge left in the
furnace) is carried out. In this process, however, the glowing material
brought out could be quenched with cold water to such a point that the
evolution of metallic vapors could be largely avoided.
Lastly, the mode of charging of the shaft furnace is also to be
regarded as a cause of poisoning, inasmuch as it is impossible to
avoid entirely the raising of dust in the repeated act of dumping and
turning over the materials for smelting, in preparing the mix, and in
subsequently charging the furnace.
By the introduction of the Huntington-Heberlein process, all these
disadvantages, both in the roasting operation and in running the shaft
furnaces, are in part removed altogether, in part reduced to such a
degree that the danger of injury is brought to a minimum.
In furnaces in which the product of the Huntington-Heberlein roast
is smelted, the slag is tapped only periodically at considerable
intervals; and, as there is less lead matte produced than formerly, the
quantity of smoke is never so great that the ventilating fan cannot
easily take care of it. There is therefore little chance of any smoke
escaping into the working-room.
As the production of work-lead, especially in the case of the large
shaft furnace, is very considerable, so that the lead continually
flows out in a big stream into the well, the hand ladling has to
be abandoned. Therefore the lead is conducted to a large reservoir
standing near the sump, and is there allowed to cool below its
volatilizing temperature. As soon as this tank is full, the lead is
tapped off and (by the aid of a swinging gutter) is cast into molds
ready for this purpose. Both the sump and the reservoir-tank are placed
under a fume-hood. The swinging gutter is covered with sheet-iron lids
while tapping, so that any lead volatilized is conveyed by the gutter
itself to a hood attached to the reservoir; thus the escape of metallic
vapors into the working space is avoided, as far as possible.
This method of pouring does not entail the same bodily exertion as the
ladling of the lead; moreover, as it requires but little time, it gives
the workmen frequent opportunity to rest.
But one of the chief advantages of the Huntington-Heberlein process
lies in the entire omission of the barring down. If the running of the
shaft furnace is conducted with any degree of care, disorders in the
working of the furnace do not occur, and one can rely on a perfectly
regular course of the smelting process day after day. No formation
of any crusts interfering with the operation of the furnace has been
recorded during any of the campaigns, which have, in each case, lasted
nearly a year.
As regards the cleaning out of the furnace, this cannot be avoided
on blowing out the Huntington-Heberlein shaft furnace; but at most
it occurs only once a year, and can be done with less danger to the
workmen, owing to the better equipment.
Further, the charge is thrown straight into the furnace (in the case
of the large shaft furnace); thus the repeated turning over of the
smelting material, as formerly practised, becomes unnecessary, and the
deleterious influence of the unavoidable formation of dust is much
diminished.
The accompanying statistics of sickness due to lead-poisoning in
connection with the operation of the shaft furnace (referring to the
same period of time as those given above for the roasting furnaces)
confirm the above statements.
────┬──────────┬────────────────────────────────────────────
│ │ LEAD-POISONING—SHAFT FURNACES
│ ├─────────────────────┬──────────────────────
YEAR│NO. OF MEN│ CASES │ DAYS OF ILLNESS
│ ├─────┬───────────────┼─────┬────────────────
│ │TOTAL│PER 100 PERSONS│TOTAL│PER 100 PERSONS
────┼──────────┼─────┼───────────────┼─────┼────────────────
1902│ 250 │ 58 │ 23.2 │ 956 │ 382.4
1903│ 267 │ 59 │ 22.1 │1044 │ 391.0
1904│ 232 │ 24 │ 10.3 │ 530 │ 228.4
────┴──────────┴─────┴───────────────┴─────┴────────────────
If it were possible to make the necessary distinctions in the case of
the large shaft furnace, the diminution in sickness from lead-poisoning
would be still more apparent; for, among the furnace attendants proper,
there has been no illness; all cases of poisoning have occurred among
the men who prepare the charge, who break up the roasted material, and
others who are occupied with subsidiary work. Some of these are exposed
to illness through their own fault, owing to want of cleanliness, or to
neglect of every precautionary measure against lead-poisoning.
Thus far we have dealt only with the advantages and improvements of the
Huntington-Heberlein process; we will now, in conclusion, consider also
its disadvantages.
The chief drawback of the new process lies in the difficulty of
breaking up the blocks of the roasted product from the converters, a
labor which, apart from the great expense involved, is also unhealthy
for the workmen engaged thereon. Seemingly this evil is still further
increased by working with larger charges in the 10 ton converters, as
projected; but in this case it is proposed to place the converters in
an elevated position, and to cause the blocks to be shattered by their
fall from a certain hight, so that further breaking up will require
but little work. Trials made in this direction have already yielded
satisfactory results, and seem to promise that the disadvantage will in
time become less important.
Another unpleasant feature is the presence (in the waste gases from the
converters) of a higher percentage of sulphur dioxide, the suppression
of which, if it is feasible at all, might be fraught with trouble and
expense.
That the roaster gases from the reverberatory-smelting and sintering
furnaces did not show such a high percentage of sulphur dioxide must
be ascribed chiefly to the circumstance that the roasting was much
slower, and that the gases were largely diluted with air already at the
point where they are formed, as the work must always be done with the
working-doors open. In the Huntington-Heberlein process, on the other
hand, the aim is to prevent, as far as possible, the access of air from
outside while blowing the charge. The more perfectly this is effected,
and the greater the quantity of ore to be blown in the converters, the
higher will also be the percentage of sulphur dioxide in the waste
gases. This circumstance has not only furnished the inducement, but it
has rendered it possible to approach the plan of utilizing the sulphur
dioxide for the manufacture of sulphuric acid. If this should be done
successfully (which, according to the experiments carried out, there
is reasonable ground to expect), the present disadvantage might be
turned into an advantage. This has the more significance because an
essential constituent of the lead ore—the sulphur—will then no longer,
as hitherto, have to be regarded as wholly lost.[31]
THE HUNTINGTON-HEBERLEIN PROCESS
BY THOMAS HUNTINGTON AND FERDINAND HEBERLEIN
(May 26, 1906)
This process for roasting lead sulphide ores has now fairly
established itself in all parts of the world, and is recognized by
metallurgical engineers as a successful new departure in the method of
desulphurization. It offers the great advantage over previous methods
of being a more scientific application of the roasting reactions (of
the old well-used formulæ PbS + 3O = PbO + SO₂ and PbS + PbSO₄
+ 2O = 2PbO + 2SO₂) and admits of larger quantities being handled
at a time, so that the use of fuel and labor are in proportion to the
results achieved, and also there is less waste all around in so far
as the factors necessary for the operation—fuel, labor and air—can
be more economically used. The workman’s time and strength are not
employed in laboriously shifting the ore from one part of the furnace
to another with a maximum amount of exertion and a minimum amount of
oxidation. The fuel consumed acts more directly upon the ore during the
first part of the process in the furnace and its place is taken by the
sulphur itself during the final and blowing stage, so that during the
whole series of operations more concentrated gases are produced and
consequently the large excess of heated air of the old processes is
avoided to such an extent that the gases can be used for the production
of sulphuric acid.
With a modern well-constructed plant practically all the evils of
the old hand-roasting furnaces are avoided, and besides the notable
economy achieved by the H.-H. process itself, the health and well-being
of the workmen employed are greatly advanced, so that where hygienic
statistics are kept it is proved that lead-poisoning has greatly
diminished. It is only natural, therefore, that the H.-H. process
should have been a success from the start, popular alike with managers
and workmen once the difficulties inseparable from the introduction of
any new process were overcome.
Simple as the process now appears, however, it is the result of many
years of study and experiment, not devoid of disappointments and at
times appearing to present a problem incapable of solution. The first
trials were made in the smelting works at Pertusola, Italy, as far
back as 1889, where considerable sums were devoted every year to this
experimental work and lead ore roasting was almost continuously on the
list of new work from 1875 on.
It may be interesting to mention that at this time the Montevecchio
ores (containing about 70 per cent. lead and about 15 per cent.
sulphur, together with a certain amount of zinc and iron) were
considered highly refractory to roast, and the only ores approved of
by the management of the works at this date were the Monteponi and
San Giovanni first-class ores (containing about 80 per cent. lead),
and the second-class carbonates (with at least 60 per cent. lead and
5 per cent. sulphur). It must be noted that a modified Flintshire
reverberatory process was in use in the works, which could deal
satisfactorily only with this class of ore, so that, as these easy ores
diminished in quantity every year and their place was taken by the
“refractory” Montevecchio type, the roasting problem was always well to
the front at the Pertusola works.
It may be asserted that almost every known method of desulphurization
was examined and experimented upon on a large scale. Gas firing was
exclusively used on certain classes of ores for several years with
considerable success, and revolving furnaces of the Brückner type—gas
fired—were also tried. Although varying degrees of success were
obtained, no really great progress was made in actual desulphurization;
methods were cheapened and larger quantities handled at a time, but
the final product—whether sintered or in a pulverulent state—seldom
averaged much under 5 per cent. sulphur, while the days of the
old “gray slags” (1 per cent. to 2 per cent. sulphur) from the
reverberatories totally disappeared, together with the class of ores
which produced them.
During the long period of these experiments in desulphurization various
facts were established:
(1) That sulphide of lead—especially in a pulverulent state—could not
be desulphurized in the same way as other sulphides, such as sulphides
of iron, copper, zinc, etc., because if roasted in a mechanical
furnace the temperature had to be kept low enough to avoid premature
sintering, which would choke the stirrers and cause trouble by the
ore clogging on the sides and bottom of the furnace. If, however, the
ore was roasted in a “dry state” at low temperature, a great deal of
sulphur remained in the product as sulphate of lead, which was as
bad for the subsequent blast-furnace work as the sulphide of lead
itself. When air was pressed through molten galena—in the same way as
through molten copper matte—a very heavy volatilization of lead took
place, while portions of it were reduced to metal or were contained as
sulphide in the molten matte, so that a good product was not obtained.
(2) That no complete dead roast of lead ores could be obtained unless
the final product was thoroughly smelted and agglomerated.
(3) That a well roasted lead ore could be obtained by oxidizing the PbS
with compressed air, after the ore had been suitably prepared.
(4) That metal losses were mainly due to the excessive heat produced in
the oxidation of PbS to PbO, and other sulphides present in the ore.
It was by making use of these facts that the H.-H. roasting process
was finally evolved, and by carefully applying its principles it is
possible to desulphurize completely the ore to a practically dead roast
of under 1 per cent. sulphur; in practice, however, such perfection
is unnecessary and a well agglomerated product with from 2 to 2.5 per
cent. sulphur is all that is required. During some trials in Australia,
where a great degree of perfection was aimed at, a block of over 2000
tons of agglomerated, roasted ore was produced containing 1 per cent.
sulphur (as sulphide); as the ores contained an average of about 10 per
cent. Zn, this was a very fine result from a desulphurization point
of view, but it was not found that this 1 per cent. product gave any
better results in the subsequent smelting in the blast furnace than
later on a less carefully prepared material containing 2.5 per cent.
sulphur.
In the early stages of experiment the great difficulty was to obtain
agglomeration without first fusing the sulphides in the ore, and
turning out a half-roasted product full of leady matte. Simple as the
thing now is, it seemed at times impossible to avoid this defect, and
it was only by a careful study of the effects of an addition of lime,
Fe₂O₃ or Mn₂O₃, and their properties that the right path
was struck. Before the introduction of the H.-H. process lime was
only used in the reverberatory process (Flintshire and Tarnowitz) to
stiffen the charge, but as Percy tells us that after its addition the
charge was glowing, it must have had a chemical as well as a mechanical
effect. In recognition of this fact fine caustic lime or crushed
limestone was mixed with the ore _before_ charging it into the furnace
and exposing it to an oxidizing heat.
It was thought probable that a dioxide of lime might be temporarily
formed, which in contact with PbS would be decomposed immediately after
its formation, or that the CaO served as _Contactsubstanz_ in the same
way as spongy platinum, metallic silver, or oxide of iron. As CaSO₄
and not CaSO₃ is always found in the roasted ore, this may prove
that CaO is really a contact substance for oxygen (see W. M. Hutchings,
_Engineering and Mining Journal_, Oct. 21, 1905, Vol. LXXX, p. 726).
The fact that the process works equally well with Fe₂O₃ instead
of CaO speaks against the theory of plumbate of lime. Whatever theory
may be correct, the fact remains that CaO assists the roasting process
and that by its use the premature agglomeration of the sulphide ore is
avoided. A further advantage of lime is that it keeps the charge more
porous and thus facilitates the passage of the air.
The shape and size of the blowing apparatus best adapted for the
purpose in view occupied many months; starting from very shallow
pans or rectangular boxes several feet square with a few inches of
material over a perforated plate, it gradually resolved itself into the
cone-shaped receptacle—holding about a ton of ore—as first introduced
together with the process. In later years and in treating larger
quantities a more hemispherical form has been adopted, containing up to
15 tons of ore.
It is probable about eight years were employed in actually working out
the process before it was introduced on any large scale at Pertusola,
but by the end of 1898 the greater part of the Pertusola ores were
treated by the process. Its first introduction to any other works was
in 1900, when it was started outside its home for the first time at
Braubach (Germany). Since then its application has gradually extended,
proceeding from Europe to Australia and Mexico and finally to America
and Canada, where recognition of its merits was more tardy than
elsewhere. It is now practically in general use all over the world and
is recognized as a sound addition to metallurgical progress. It is
doubtless only a step in the right direction and with its general use
a better knowledge of its principles will prevail, so that its future
development in one direction or another, as compared with present
results, may show some further progress.
The present working of the H.-H. process still follows practically the
original lines laid down, and by preliminary roasting in a furnace
with lime, oxide of iron, or manganese (if not already contained in
the ore), prepares the ore for blowing in the converter. Mechanical
furnaces have been introduced to the entire exclusion of the old
hand-roasters, and the size of the converters has been gradually
increased from the original one-ton apparatus successively to 5, 7
and 10 ton converters; at present some for 15 tons are being built in
Germany and will doubtless lead to a further economy.
The mechanical furnace at present most in use is a single-hearth
revolving furnace with fixed rabbles, the latest being built with a
diameter of 26½ ft. and a relatively high arch to ensure a clear flame
and rapid oxidation of the ore. The capacity of these furnaces varies,
of course, with the nature of the ores to be treated, but with ordinary
lead ores (European and Australian practice) of from 50 per cent. to 60
per cent. lead and 14 per cent, to 18 per cent. sulphur, the average
capacity may be taken at about 50 to 60 tons of crude ore per day of
24 hours. The consumption of coal with a well-constructed furnace is
very low and is always under 8 per cent.—6 per cent. being perhaps the
average. These furnaces require very little attention, being automatic
in their charging and discharging arrangements.
The ore on leaving the furnace is charged into the converters by
various mechanical means (Jacob’s ladders, conveyors, etc.). The
converter charge usually consists of some hot ore direct from the
furnace, on top of which ore is placed which has been cooled down by
storage in bins or by the addition of water. The converter is generally
filled in two charges of five tons each, and the blowing time should
not be more than 4 to 6 hours. The product obtained should be porous
and well agglomerated, but easily broken up, tough melted material
being due to an excess of silica and too much lead sulphide. Attention,
therefore, to these two points (good preliminary roasting and
correction of the charge by lime) obviates this trouble. This roasted
ore should not contain more than about 1.5 to 2 per cent. sulphur,
and in a modern blast furnace gives surprisingly good results, the
matte-fall being in most cases reduced to nothing, and the capacity of
the furnace is largely increased, while the slags are poorer.
If the converter charge has been properly prepared, the blowing
operation proceeds with the greatest smoothness and requires very
little attention on the part of the workmen, the heat and oxidation
rise gradually from the bottom and volatilization losses remain low, so
that it is possible, if desired, to produce hot concentrated sulphurous
gases suitable for the manufacture of sulphuric acid.
Besides the actual economy obtained in roasting ores by the process,
a great feature of its success has been the remarkable improvement
in smelting and reducing the roasted ore as compared with previous
experience. This is due to the nature of the roasted material, which,
besides being much poorer in sulphur than was formerly the case, is
thoroughly porous and well agglomerated and contains—if the original
mixture is properly made—all the necessary slagging materials itself,
so that it practically becomes a case of smelting slags instead of ore,
and to an expert the difference is evident.
Experience has shown that on an average the improvement in the capacity
of the blast furnace may be taken at about 50 to 100 per cent., so that
in works using the H.-H. process—after its complete introduction—about
half the blast furnaces formerly necessary for the same tonnage were
blown out. The matte-fall has become a thing of the past, so that,
except in those cases where some matte is required to collect the
copper contained in the ores, lead matte has disappeared and the
quantity of flue dust as well as the lead and silver losses have been
greatly reduced.
Referring to the latest history of the H.-H. process, and the theory
of direct blowing, it may be remarked—putting aside all legal
questions—that the idea, metallurgically speaking, is attractive, as it
would seem that by eliminating one-half of the process and blowing the
ores direct without the expense of a preliminary roast a considerable
economy should be effected. Upon examination, however, this supposed
economy and simplicity is not at all of such great importance, and
in many cases, without doubt, would be retrogressive in lead ore
smelting rather than progressive. When costs of roasting in a furnace
are reduced to such a low figure as can be obtained by using 50 ton
furnaces and 10 to 15 ton converters, there is very little margin
for improvement in this direction. From the published accounts of
the Tarnowitz smelting works (the _Engineering and Mining Journal_,
Sept. 23, 1905, Vol. LXXX, p. 535) the cost of mechanical preliminary
roasting cannot exceed 25c. per ton, so that even assuming direct
blowing were as cheap as blowing a properly prepared material, the
total economy would only be the above figure, viz., 25c.; but this is
far from being the case.
Direct blowing of a crude ore is considerably more expensive than
dealing with the H.-H. product, because of necessity the blowing
operation must be carried out slowly and with great care so as to avoid
heavy metal losses, and whereas a pre-roasted ore can be easily blown
in four hours and one man can attend to two or three 10 ton converters,
the direct blowing operation takes from 12 to 18 hours and requires the
continual attention of one man. In the first case the cost of labor
would be: One man at say $3 for 50 tons (at least), i.e., 6c. per
ton, and in the second case one man at $3 for 10 tons (at the best),
i.e., 30c., a difference in favor of pre-roasting of 24c., so that any
possible economy would disappear. Furthermore, as the danger of blowing
upon crude sulphides for 12 or 18 hours is greater as regards metal
losses than a quick operation of four hours, it is very likely that
instead of an economy there would be an increase in cost, owing to a
greater volatilization of metals.
These remarks refer to ordinary lead ores with say 50 per cent. lead
and about 14 per cent. sulphur. With ores, however, such as are
generally treated in the United States the advantages of pre-roasting
are still more evident. These ores contain about 10 to 15 per cent.
lead, 30 to 40 per cent. sulphur, 20 to 30 per cent. iron, 10 per cent.
zinc, 5 per cent. silica, and lose the greater part of the pyritic
sulphur in the preliminary roasting, leaving the iron in the form of
oxide, which in the subsequent blowing operation acts in the same
way as lime. For this reason the addition of extra fluxes, such as
limestone, gypsum, etc., to the original ore is not necessary and only
a useless expense.
In certain exceptional cases and with ores poor in sulphur, direct
blowing might be applicable, but for the general run of lead ores no
economy can be expected by doing away with the preliminary roast.
MAKING SULPHURIC ACID AT BROKEN HILL
(August 11, 1904)
The Broken Hill Proprietary Company has entered upon the
manufacture of sulphuric acid on a commercial scale. The acid is
practically a by-product, being made from the gases emanating
from the desulphurization of the ores, concentrates, etc., by the
Carmichael-Bradford process. The acid can be made at a minimum of
cost, and thus materially enhances the value of the process recently
introduced for the separation of zinc blende from the tailings by
flotation. The following particulars are taken from a recently
published description of the process: The ores, concentrates, slimes,
etc., as the case may be, are mixed with gypsum, the quantity of the
latter varying from 15 to 25 per cent. The mixture is then granulated
to the size of marbles and dumped into a converter. The bottom of
the charge is heated from 400 to 500 deg. C. It is then subjected to
an induced current of air, and the auxiliary heat is turned off. The
desulphurization proceeds very rapidly with the evolution of heat and
the gases containing sulphurous anhydride. The desulphurization is very
thorough, and no losses occur through volatilization. The sulphur thus
rendered available for acid making is rather more than is contained in
the ore, the sulphur in the agglomerated product being somewhat less
than that accounted for by the sulphur contained in the added gypsum.
Thus from one ton of 14 per cent. sulphide ore it is possible to make
about 12 cwt. of chamber acid, fully equaling 7 cwt. of strong acid.
The plant at present in use, which comprises a lead chamber of 40,000
cu. ft., can turn out 35 tons of chamber acid per week. This plant is
being duplicated, and it has also been decided to erect a large plant
at Port Pirie for use in the manufacture of superphosphates. It is
claimed that the production of sulphuric acid from ores containing only
14 per cent. of sulphur establishes a new record.
THE CARMICHAEL-BRADFORD PROCESS
BY DONALD CLARK
(November 3, 1904)
Subsequent to the introduction of the Huntington-Heberlein process
in Australia, Messrs. Carmichael and Bradford, two employees of the
Broken Hill Proprietary Company, patented a process which bears their
name. Instead of starting with lime, or limestone and galena, as in
the Huntington-Heberlein process, they discovered that if sulphate of
lime is mixed with galena and the temperature raised, on blowing a
current of air through the mixture the temperature rises and the mass
is desulphurized. The process would thus appear to be a corollary of
the original one, and the reactions in the converter are identical.
Owing to the success of the acid processes in separating zinc sulphide
from the tailing at Broken Hill, it became necessary to manufacture
sulphuric acid locally in large quantity. The Carmichael-Bradford
process has been started for the purpose of generating the sulphur
dioxide necessary, and is of much interest as showing how gases rich
enough in SO₂ may be produced from a mixture containing only from 13
to 16 per cent. sulphur.
Gypsum is obtained in a friable state within about five miles from
Broken Hill. This is dehydrated, the CaSO, 2H₂O being converted into
CaSO₄ on heating to about 200 deg. C. The powdered residue is mixed
with slime produced in the milling operations and concentrate in the
proportion of slime 3 parts, concentrate 1 part, and lime sulphate 1
part. The proportions may vary to some extent, but the sulphur contents
run from 13 to 16 or 17 per cent. The average composition of the
ingredients is as given in the table on the next page.
These materials are moistened with water and well mixed by passing
them through a pug-mill. The small amount of water used serves to
set the product, the lime sulphate partly becoming plaster of paris,
2CaSO, H₂O. While still moist the mixture is broken into pieces not
exceeding two inches in diameter and spread out on a drying floor,
where excess of moisture is evaporated by the conjoint action of sun
and wind.
─────────────────┬─────┬───────────┬────────┬────────
│SLIME│CONCENTRATE│CALCIUM │AVERAGE
│ │ │SULPHATE│
──────────────────┼─────┼───────────┼────────┼────────
Galena │ 24 │ 70 │ │ 29
Blende │ 30 │ 15 │ │ 21
Pyrite │ 3 │ │ │ 2
Ferric oxide │ 4 │ │ │ 2.5
Ferrous oxide │ 1 │ │ │ 1
Manganous oxide │ 6.5│ │ │ 5
Alumina │ 5.5│ │ │ 3
Lime │ 3.5│ │ 41 │ 10
Silica │ 23 │ │ │ 14
Sulphur trioxide │ │ │ 59 │ 12
─────────────────┴─────┴───────────┴────────┴────────
The pots used are small conical cast-iron ones, hung on trunnions,
and of the same pattern as used in the Huntington-Heberlein process.
Three of these are set in line, and two are at work while the third is
being filled. These pots have the same form of conical cover leading
to a telescopic tube, and all are connected to the same horizontal
pipe leading to the niter pots. Dampers are provided in each case. A
small amount of coal or fuel is fed into the pots and ignited by a
gentle blast; as soon as a temperature of about 400 to 500 deg. C. is
attained the dried mixture is fed in, until the pot is full; the cover
is closed down and the mass warms up. Water is first driven off, but
after a short time concentrated fumes of sulphur dioxide are evolved.
The amount of this gas may be as much as 14 per cent., but it is
usually kept at about 10 per cent., so as to have enough oxygen for
the conversion of the dioxide to the trioxide. The gases are led over
a couple of niter pots and thence to the usual type of lead chamber
having a capacity of 40,000 cu. ft. Chamber acid alone is made, since
this requires to be diluted for what is known as the saltcake process.
The plant has now been in operation for some time and, it is claimed,
with highly successful results. The product tipped out of. the
converter is similar to that obtained in the Huntington-Heberlein
process, and is at once fit for the smelters, the amount of sulphur
left in it being always less than that originally introduced with the
gypsum; analysis of the desulphurized material shows usually from 3 to
4 per cent. sulphur.
THE CARMICHAEL-BRADFORD PROCESS
BY WALTER RENTON INGALLS
(October 28, 1905)
As described in United States patent No. 705,904, issued July 29, 1902,
lead sulphide ore is mixed with 10 to 35 per cent. of calcium sulphate,
the percentage varying according to the grade of the ore. The mixture
is charged into a converter and gradually heated externally until the
lower portion of the charge, say one-third to one-fourth, is raised to
a dull-red heat; or the reactions may be started by throwing into the
empty converter a shovelful of glowing coal and turning on a blast of
air sufficient to keep the coal burning and then feeding the charge
on top of the coal. This heating effects a reaction whereby the lead
sulphide of the ore is oxidized to sulphate and the calcium sulphate is
reduced to sulphide. The heated mixture being continuously subjected
to the blast of air, the calcium sulphide is re-oxidized to sulphate
and is thus regenerated for further use. This reaction is exothermic,
and sufficient heat is developed to complete the desulphurization of
the charge of ore by the concurrent reactions between the lead sulphate
(produced by the calcium sulphate) and portions of undecomposed ore,
sulphurous anhydride being thus evolved. The various reactions, which
are complicated in their nature, continue until the temperature of
the charge reaches a maximum, by which time the charge has shrunk
considerably in volume and has a tendency to become pasty. This becomes
more marked as the production of lead oxide increases, and as the
desired point of desulphurization is attained the mixture fuses; at
this stage the calcium sulphide which is produced from the sulphate
cannot readily oxidize, owing to the difficulty of coming into actual
contact with the air in the pasty mass, but, being subjected to the
strong oxidizing effect of the metallic oxide, it is converted into
calcium plumbate, while sulphurous anhydride is set free. The mass then
cools, as the exothermic reactions cease, and can be readily removed to
a blast furnace for smelting.
The reactions above described are as outlined in the original
American patent specification. Irrespective of their accuracy,
the Carmichael-Bradford process is obviously quite similar to the
Huntington-Heberlein, and doubtless owes its origin to the latter. The
difference between them is that in the Huntington-Heberlein process
the ore is first partially roasted with addition of lime, and is then
converted in a special vessel. In the Carmichael-Bradford process
the ore is mixed with gypsum and is then converted directly. The
greatest claim for originality in the Carmichael-Bradford process
may be considered to lie in it as a method of desulphurizing gypsum,
inasmuch as not only is the sulphur of the ore expelled, but also a
part of the sulphur of the gypsum; and the sulphur is driven off as a
gas of sufficiently high tenor of sulphur dioxide to enable sulphuric
acid to be made from it economically. Up to the present time the
Carmichael-Bradford process has been put into practical use only at
Broken Hill, N. S. W.
The Broken Hill Proprietary Company first conducted a series of tests
in a converter capable of treating a charge of 20 cwt. These tests were
made at the smelting works at Port Pirie. Exhaustive experiments made
on various classes of ores satisfactorily proved the general efficacy
of the process. The following ores were tried in these preliminary
experiments, viz.:
First-grade concentrate containing: Pb, 60 per cent.; Zn, 10 per cent.;
S, 16 per cent.; Ag, 30 oz.
Second-grade concentrate containing: Pb, 45 per cent.; Zn, 12.5 per
cent.; S, 14.5 per cent.; Ag, 22 oz.
Slime containing: Pb, 21 per cent.; Zn, 17 per cent.; S, 13 per cent.;
Ag, 18 oz.
Lead-copper matte containing: Fe, 42 per cent.; Pb, 17 per cent.; Zn,
13.3 per cent.; Cu, 2.4 per cent.; S, 23 per cent.; Ag, 25 oz.
Other mattes, of varying composition up to 45 per cent. Pb and 100 oz.
Ag, were also tried.
The results from these preliminary tests were so gratifying that a
further set of tests was made on lead-zinc slime, with a view of
ascertaining whether any volatilization losses occurred during the
desulphurization. This particular material was chosen because of its
accumulation in large proportions at the mine, and the unsatisfactory
result of the heap roasting which has recently been practised. The
heap roasting, although affording a product containing only 7 per cent.
S, which is delivered in lump form and therefore quite suitable for
smelting, resulted in a high loss of metal by volatilization (17 per
cent. Pb, 5 per cent. Ag).
The result of nine charges of the slime treated by the
Carmichael-Bradford process was as follows:
─────────────────┬──────┬─────────────────────┬───────────────────────
│ │ ASSAYS │ CONTENTS
│ Cwt. ├────┬──────┬────┬────┼─────┬─────┬────┬──────
│ │Pb% │Ag oz.│Zn% │ S% │ Pb │ Ag. │ Zn │ S
│ │ │ │ │ │cwt. │ oz. │cwt.│cwt.
├──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Raw slime │128.1 │21.3│ 18.0 │16.8│13.1│27.28│115.3│26.2│16.78
Raw gypsum │ 54.9 │ │ │ │ │ │ │ │ 9.88
├──────┤ │ │ │ ├─────┼─────┼────┼──────
Total │183.0 │ │ │ │ │27.28│115.3│25.2│26.66
──────────────────┼──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Sintered material│109.88│20.7│ 17.2 │ │4.80│22.74│ 94.5│ │ 5.27
Middling │ 14.47│17.7│ 15.7 │ │6.20│ 2.56│ 11.3│ │ 0.89
Fines │ 11.12│19.0│ 14.8 │ │7.50│ 2.11│ 8.2│ │ 0.83
├──────┼────┼──────┼────┼────┼─────┼─────┼────┼──────
Total │135.47│ │ │ │5.17│27.41│113.0│ │ 6.99
─────────────────┴──────┴────┴──────┴────┴────┴─────┴─────┴────┴──────
These results indicated practically no volatilization of lead and
silver during the treatment, the lead showing a slight increase, viz.,
0.47 per cent., and the silver 1.13 per cent. loss. A desulphurization
of 70.4 per cent. was effected. A higher desulphurization could have
been effected had this been desired. In the above tabulated results,
the term “middling” is applied to the loose fritted lumps lying on the
top of the charge: these are suitable for smelting, the fines being the
only portion which has to be returned.
In order to test the practicability of making sulphuric acid, a plant
consisting of three large converters of capacity of five tons each,
together with a lead chamber 100 ft. by 20 ft. by 20 ft., was then
erected at Broken Hill, together with a dehydrating furnace, pug-mill,
and granulator. These converters are shown in the accompanying
engravings.
A trial run was made with 108 tons of concentrate of the following
composition: 54 per cent. lead; 1.9 per cent. iron; 0.9 per cent.
manganese; 9.4 per cent. zinc; 14.6 per cent. sulphur; 19.2 per cent.
insoluble residue, and 24 oz. silver per ton.
The converter charge consisted of 100 parts of the concentrate and
25 parts of raw gypsum, crushed to pass a 1 in. hole and retained
by a 0.25 in. hole, the material finer than 0.25 in. (which amounted
to 5 per cent. of the total) being returned to the pug-mill. After
desulphurization in the converter, the product assayed as follows:
48.9 per cent. lead; 1.80 per cent. iron; 0.80 per cent. manganese;
7.87 per cent. zinc; 3.90 per cent. sulphur; 1.02 per cent. alumina;
5.80 per cent. lime; 21.75 per cent. insoluble residue; 8.16 per cent.
undetermined (oxygen as oxides, sulphates, etc.); total, 100 per cent.
Its silver content was 22 oz. per ton. The desulphurized ore weighed
10 per cent. more than the raw concentrate. During this run 34 tons of
acid were made.
A trial was then made on 75 tons of slime of the following composition:
18.0 per cent. lead; 16.6 per cent. zinc; 6.0 per cent. iron; 2.5 per
cent. manganese; 3.2 per cent. alumina; 2.1 per cent. lime; 38.5 per
cent. insoluble residue; total, 100 per cent. Its silver content was
19.2 oz. per ton.
The converter charge in this case consisted of 100 parts of raw slime
and 30 parts of gypsum. The converted material assayed as follows:
16.1 per cent. lead; 14.0 per cent. zinc; 3.6 per cent. sulphur; 5.42
per cent. iron; 2.25 per cent. manganese; 4.10 per cent. alumina; 8.60
per cent. lime; 39.80 per cent. insoluble residue; 6.13 per cent.
undetermined (oxygen, etc.); total, 100 per cent.; and silver, 17.5
oz. per ton. The increase in weight of desulphurized ore over that
of the raw ore was 11 per cent. During this run 22 tons of acid were
manufactured.
The analysis of the gypsum used in each of the above tests (at Broken
Hill) was as follows: 76.1 per cent. CaSO₄, 2H₂O; 0.5 per cent.
Fe₂O₃; 4.5 per cent. Al₂O₃; 18.9 per cent. insoluble
residue.
The plant was then put into continuous operation on a mixture of three
parts slime and one of concentrate, desulphurizing down to 4 per cent.
S, and supplying 20 tons of acid per week, and additions were made to
the plant as soon as possible. The acid made at Broken Hill has been
used in connection with the Delprat process for the concentration of
the zinc tailing. At Port Pirie, works are being erected with capacity
for desulphurization of about 35,000 tons per annum, with an acid
output of 10,000 tons. This acid is to be utilized for the acidulation
of phosphate rock.
[Illustration: FIG. 15.—Details of Converters.]
The cost of desulphurization of a ton of galena concentrate by the
Carmichael-Bradford process, based on labor at $1.80 per 8 hours,
gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb., is
estimated as follows:
0.25 ton of gypsum $0.60
Dehydrating and granulating gypsum .48
Drying mixture of ore and gypsum .12
Converting .24
Spalling sintered material .12
0.01 ton coal .08
——-——-
Total $1.64
The lime in the sintered product is credited at 12c., making the net
cost $1.52 per ton (2240 lb.) of ore.
The plant required for the Carmichael-Bradford process can be described
with sufficient clearness without drawings, except the converters. The
ore (concentrate, slime, etc.) to be desulphurized is delivered at the
top of the mill by cars, conveyors, or other convenient means, and
dumped into a bin. Two screw feeders placed inside the bin supply the
mill with ore, uniformly and as fast as it is required. These feeders
deliver the ore into a chute, which directs it into a vertical dry
mixer.
A small bin, on the same level as the ore-bin, receives the crude
gypsum from cars. Thence it is fed automatically to a disintegrator,
which pulverizes it finely and delivers it into a storage bin
underneath. This disintegrator revolves at about 1700 r.p.m. and
requires 10 h.p. The body of the machine is cast iron, fitted with
renewable wearing plates (made of hard iron) in the grinding chamber.
The revolving parts consist of a malleable iron disc in which are fixed
steel beaters, faced on the grinding surface with highly tempered
steel. The bin that receives the floured gypsum contains a screw
conveyor similar to those in the ore-bin, and dumps the material into
push conveyors passing into the dehydrating furnace. They carry the
crushed gypsum along at a speed of about 1 ft. per minute, and allow
about 20 ft. to dehydrate the gypsum. This speed can, of course, be
regulated to suit requirements.
The dehydrated gypsum runs down a chute into an elevator boot, and is
elevated into a bin which is on the same level as the ore-bin. This bin
also contains a screw conveyor, like that in the ore-bin. The speed of
delivery is regulated to deliver the right proportion of dehydrated
gypsum to the mixer.
The mixer is of the vertical pattern and receives the sulphide ore
and dehydrated gypsum from the screw feeders. In it are set two flat
revolving cones running at different speeds, thus ensuring a thorough
mixture of the gypsum and ore. The mixed material drops from the
cones upon two baffle plates, and is wetted just before entering the
pug-mill. The pug-mill is a wrought-iron cylinder of ¼ in. plate about
2 ft. 6 in. diameter and 6 or 8 ft. long, and has the mixer fitted
to the head. It contains a 3 ft. wrought-iron spiral with propelling
blades, which forces the plastic mixture through ¾ in. holes in the
cover. The material comes out in long cylindrical pieces, but is broken
up and formed into marble-shaped pieces on dropping into a revolving
trommel.
The trommel is about 5 ft. long, 2 ft. in diameter at the small end and
about 4 ft. at the large end. It revolves about a wrought-iron spindle
(2½ in. diameter) carrying two cast-iron hubs to which are fitted
arms for carrying the conical plate ⅛ in. thick. About 18 in. of
the small end of the cone is fitted with wire gauze, so as to prevent
the material as it comes out of the pug-mill from sticking to it. The
trommel is driven by bevel gearing at 20 to 25 r.p.m. The granulated
material formed in the trommel is delivered upon a drying conveyor.
The conveyor consists of hinged wrought-iron plates flanged at the side
to keep the material from running off. It is driven from the head by
gearing, at a speed of 1 ft. per minute, passing through a furnace 10
ft. long to dry and set the granules of ore and gypsum. This speed can,
of course, be regulated to suit requirements. The granulated material,
after leaving the furnace, is delivered to a single-chain elevator,
traveling at a speed of about 150 ft. per minute. It drops the material
into a grasshopper conveyor, driven by an eccentric, which distributes
the material over the length of a storage bin. From this bin the
material is directed into the converters by means of the chutes, which
have their bottom ends hinged so as to allow for the raising of the
hood when charging the converters.
The converters are shown in the accompanying engravings, but they may
be of slightly different form from what is shown therein, i.e., they
may be more spherical than conical. They will have a capacity of about
four tons, being 6 ft. in diameter at the top, 4 ft. in diameter at
the false bottom, and about 5 ft. deep. They are swung on cast-iron
trunnions bolted to the body and turned by means of a hand-wheel and
worm (not shown). They are carried on strong cast-iron standards fitted
with bearings for trunnions, and all necessary brackets for tilting
gear. The hood has a telescopic funnel which allows it to be raised
or lowered, weights being used to balance it. At the apex of the cone
a damper is provided to regulate the draft. A 4 in. hole in the pot
allows the air from the blast-pipe, 18 in. in diameter, to enter under
the false perforated bottom, the connection between the two being made
by a flexible pipe and coupling. Two Baker blowers supply the blast for
the converters. The material, after being sintered, is tipped on the
floor in front of the converters and is there broken up to any suitable
size, and thence dispatched to the smelters.
[Illustration: FIG. 16.—Arrangement of Converters.]
The necessary power for a plant with a capacity of 150 tons of ore per
day will be supplied by a 50 h.p. engine.
THE SAVELSBERG PROCESS
BY WALTER RENTON INGALLS
(December 9, 1905)
There are in use at the present time three processes for the
desulphurization of galena by the new method, which has been referred
to as the “lime-roasting of galena.” The Huntington-Heberlein and the
Carmichael-Bradford processes have been previously described. The third
process of this type, which in some respects is more remarkable than
either of the others, is the invention of Adolf Savelsberg, director
of the smeltery at Ramsbeck, Westphalia, Germany, which is owned by
the Akt. Gesell. f. Bergbau, Blei. u. Zinkhüttenbetrieb zu Stolberg
u. in Westphalen. The process is in use at the Ramsbeck and Stolberg
lead smelteries of that company. It is described in American patent
No. 755,598, issued March 22, 1904 (application filed Dec. 18, 1903).
The process is well outlined in the words of the inventor in the
specification of that patent:
“The desulphurizing of certain ores has been effected by blowing air
through the ore in a chamber, with the object of doing away with the
imperfect and costly process of roasting in ordinary furnaces; but
hitherto it has not been possible satisfactorily to desulphurize lead
ores in this manner, as, if air be blown through raw lead ores in
accordance with either of the processes used for treating copper ores,
for example, the temperature rises so rapidly that the unroasted lead
ore melts and the air can no longer act properly upon it, because
by reason of this melting the surface of the ores is considerably
decreased, the greater number of points or extent of surface which
the raw ore originally presented to the action of the oxygen of the
air blown through being lost, and, moreover, the further blowing
of air through the molten mass of ore produces metallic lead and a
plumbiferous slag (in which the lead oxide combines with the gangue)
and also a large amount of light dust, consisting mainly of sublimated
lead sulphide. Huntington and Heberlein have proposed to overcome
these objections by adopting a middle course, consisting in roasting
the ores with the addition of limestone for overcoming the ready
fusibility of the ores, and then subjecting them to the action of the
current of air in the chamber; but this process is not satisfactory,
because it still requires the costly previous operation in a roasting
furnace.
[Illustration: Fig. 18.—Converter Ready to Dump.]
“My invention is based on the observation which I have made that if
the lead ores to be desulphurized contain a sufficient quantity of
limestone it is possible, by observing certain precautions, to dispense
entirely with the previous roasting in a roasting furnace, and to
desulphurize the ores in one operation by blowing air through them. I
have found that the addition of limestone renders the roasting of the
lead ore unnecessary, because the limestone produces the following
effects:
“The particles of limestone act mechanically by separating the
particles of lead ore from each other in such a way that premature
agglomeration is prevented and the whole mass is loosened and rendered
accessible to air; and, moreover, the limestone moderates the high
reaction temperature resulting from the burning of the sulphur, so
that the liquefaction of the galena, the sublimation of lead sulphide,
and the separation of metallic lead are avoided. The moderation of
the temperature of reaction is caused by the decomposition of the
limestone into caustic lime and carbon dioxide, whereby a large amount
of heat becomes latent. Further, the decomposition of the limestone
causes chemical reactions, lime being formed, which at the moment of
its formation is partly converted into sulphate of lime at the expense
of the sulphur contained in the ore, and this sulphate of lime, when
the scorification takes place, is transformed into calcium silicate
by the silicic acid, the sulphuric acid produced thereby escaping.
The limestone also largely contributes to the desulphurization of the
ore, as it causes the production of sulphuric acid at the expense of
the sulphur of the ore, which sulphuric acid is a powerful oxidizing
agent. If, therefore, a mixture of raw lead ore and limestone (which
mixture must, of course, contain a sufficient amount of silicic acid
for forming silicates) be introduced into a chamber and a current of
air be blown through the mixture, and at the same time the part of the
mixture which is near the blast inlet be ignited, the combustion of the
sulphur will give rise to very energetic reactions, and sulphurous
acid, sulphuric acid, lead oxide, sulphates and silicates are produced.
The sulphurous acid and the carbon dioxide escape, while the sulphuric
acid and sulphates act in their turn as oxidizing agents on the
undecomposed galena. Part of the sulphates is decomposed by the silicic
acid, thereby liberating sulphuric acid, which, as already stated, acts
as an oxidizing agent. The remaining lead oxide combines finally with
the gangue of the ore and the non-volatile constituents of the flux
(the limestone) to form the required slag. These several reactions
commence at the blast inlet at the bottom of the chamber, and extend
gradually toward the upper portion of the charge of ore and limestone.
Liquefaction of the ores does not take place, for although a slag is
formed it is at once solidified by the blowing in of the air, the
passages formed thereby in the hardening slag allowing of the continued
passage therethrough of the air. The final product is a silicate
consisting of lead oxide, lime, silicic acid, and other constituents of
the ore, which now contains but little or no sulphur and constitutes a
coherent solid mass, which, when broken into pieces, forms a material
suitable to be smelted.
“The quantity of limestone required for the treatment of the lead
ores varies according to the constitution of the ores. It should,
however, amount generally to from 15 to 20 per cent. As lead ores do
not contain the necessary amount of limestone as a natural constituent,
a considerable amount of limestone must be added to them, and this
addition may be made either during the dressing of the ores or
subsequently.
“For the satisfactory working of the process, the following precautions
are to be observed: In order that the blowing in of the air may not
cause particles of limestone to escape in the form of dust before
the reaction begins, it is necessary to add to the charge before it
is subjected to the action in the chamber a considerable amount of
water—say 5 per cent. or more. This water prevents the escape of dust,
and it also contributes considerably to the formation of sulphuric
acid, which, by its oxidizing action, promotes the reaction, and,
consequently, also the desulphurization. It is advisable, in conducting
the operation, not to fill the chamber with the charge at once, but
first only partly to fill it and add to the charge gradually while the
chamber is at work, as by this means the reaction will take place more
smoothly in the mass.
[Illustration: Fig. 19.—Charge Dumped.]
“It is advantageous to proceed as follows: The bottom part of a
chamber of any suitable form is provided with a grate, on which is
laid and ignited a mixture of fuel (coal, coke, or the like) and
pieces of limestone. By mixing the fuel with pieces of limestone the
heating power of the fuel is reduced and the grate is protected,
while at the same time premature melting of the lower part of the
charge is prevented; or the grate may be first covered with a layer
of limestone and the fuel be laid thereon, and then another layer of
limestone be placed on the fuel. On the material thus placed in the
chamber, a uniform charge of lead ore and limestone—say about 12 in.
high—is placed, this having been moistened as previously explained.
Under the influence of the air-blast and the heat, the reactions
hereinbefore described take place. When the upper surface of the first
layer becomes red-hot, a further charge is laid thereon, and further
charges are gradually introduced as the surface of the preceding
charge becomes red-hot, until the chamber is full. So long as charges
are still introduced a blast of air of but low pressure is blown
through; but when the chamber is filled a larger quantity of air at a
higher pressure is blown through. The scorification process then takes
place, a very powerful desulphurization having preceded it. During the
scorification the desulphurization is completed.
“When the process is completed, the chamber is tilted and the
desulphurized mass falls out and is broken into small pieces for
smelting.”
The drawing on page 190, Fig. 17, shows a side view of the apparatus
used in connection with the process, which will be readily understood
without special description. The dotted lines show the pot in its
emptying position. The series of operations is clearly illustrated in
Figs. 18-20, which are reproduced from photographs.
This process has now been in practical use at Ramsbeck for three years,
where it is employed for the desulphurization of galena of high grade
in lead, with which are mixed quartzose silver ore (or sand if no such
ore be available), and calcareous and ferruginous fluxes. A typical
charge is 100 parts of lead ore, 10 parts of quartzose silver ore,
10 parts of spathic iron ore, and 19 parts of limestone. A thorough
mixture of the components is essential; after the mixture has been
effected, the charge is thoroughly wetted with about 5 per cent.
of water, which is conceived to play a threefold function in the
desulphurizing operation, namely: (1) preservation of the homogeneity
of the mixture during the blowing; (2) reduction of temperature during
the process; and (3) formation of sulphuric acid in the process, which
promotes the desulphurization of the ore.
[Illustration: FIG. 17.—Savelsberg Converter.]
The moistened charge is conveyed to the converters, into which it
is fed in thin layers. The converters are hemispherical cast-iron
pots, supported by trunnions on a truck, as shown in the accompanying
engravings. Except for this method of support, which renders the
pot movable, the arrangement is quite similar to that which is
employed in the Huntington-Heberlein process. The pots which are now
in use at Ramsbeck have capacity for about 8000 kg. of charge, but
it is the intention of the management to increase the capacity to
10,000 or 12,000 kg. Previously, pots of only 5000 kg. capacity were
employed. Such a pot weighed 1300 kg., exclusive of the truck. The
air-blast was about 7 cu. m. (247.2 cu. ft.) per min., beginning at
a pressure of 10 to 20 cm. of water (2¾ to 4½ oz.) and rising to
50 to 60 cm. (11½ to 13½ oz.) when the pot was completely filled with
charge. The desulphurization of a charge is completed in 18 hours. A
pot is attended by one man per shift of 12 hours; this is only the
attention of the pot proper, the labor of conveying material to it and
breaking up the desulphurized product being extra. One man per shift
should be able to attend to two pots, which is the practice in the
Huntington-Heberlein plants.
[Illustration: Fig. 20.—Converter in Position for Blowing.]
When the operation in the pot is completed, the latter is turned on its
trunnions, until the charge slides out by gravity, which it does as a
solid cake. This is caused to fall upon a vertical bar, which breaks
it into large pieces. By wedging and sledging these are reduced to
lumps of suitable size for the blast furnace. When the operation has
been properly conducted the charge is reduced to about 2 or 3 per cent.
sulphur. It is expected that the use of larger converters will show
even more favorable results in this particular.
As in the Huntington-Heberlein and Carmichael-Bradford processes, one
of the greatest advantages of the Savelsberg process is the ability to
effect a technically high degree of desulphurization with only a slight
loss of lead and silver, which is of course due to the perfect control
of the temperature in the process. The precise loss of lead has not yet
been determined, but in the desulphurization of galena containing 60
to 78 per cent. lead, the loss of lead is probably not more than 1 per
cent. There appears to be no loss of silver.
The process is applicable to a wide variety of lead-sulphide ores. The
ore treated at Ramsbeck contains 60 to 78 per cent. lead and about
15 per cent. of sulphur, but ore from Broken Hill, New South Wales,
containing 10 per cent. of zinc has also been treated. A zinc content
up to 7 or 8 per cent. in the ore is no drawback, but ores carrying a
higher percentage of zinc require a larger addition of silica and about
5 per cent. of iron ore in order to increase the fusibility of the
charge. The charge ordinarily treated at Ramsbeck is made to contain
about 11 per cent. of silica. The presence of pyrites in the ore is
favorable to the desulphurization. Dolomite plays the same part in
the process that limestone does, but is of course less desirable, in
view of the subsequent smelting in the blast furnace. The ore is best
crushed to about 3 mm. size, but good results have been obtained with
ore coarser in size than that. However, the proper size is somewhat
dependent upon the character of the ore. The blast pressure required in
the converter is also, of course, somewhat dependent upon the porosity
of the charge. Fine slimes are worked up by mixture with coarser ore.
In making up the charge, the proportion of limestone is not varied
much, but the proportions of silica and iron must be carefully modified
to suit the ore. Certain kinds of ore have a tendency to remain
pulverulent, or to retain balls of unsintered, powdered material.
In such cases it is necessary to provide more fusible material in
the charge, which is done by varying the proportions of silica and
iron. The charge must, moreover, be prepared in such a manner that
overheating, and consequently the troublesome fusion of raw galena,
will be avoided.
The essential difference between the Huntington-Heberlein and
Savelsberg processes is the use in the former of a partially
desulphurized ore, containing lime and sulphate of lime; and the use
in the latter of raw ore and carbonate of lime. It is claimed that the
latter, which loses its carbon dioxide in the converter, necessarily
plays a different chemical part from that of quicklime or gypsum.
Irrespective of the reactions, however, the Savelsberg process has the
great economic advantage of dispensing with the preliminary roasting of
the Huntington-Heberlein process, wherefore it is cheaper both in first
cost of plant and in operation.
THE LIME-ROASTING OF GALENA[32]
BY WALTER RENTON INGALLS
During the last two years, and especially during the last six
months, a number of important articles upon the new methods for the
desulphurization of galena have been published in the technical
periodicals, particularly in the _Engineering and Mining Journal_
and in _Metallurgie_. I proposed for these methods the type-name
of “lime-roasting of galena,” as a convenient metallurgical
classification,[33] and this term has found some acceptance. The
articles referred to have shown the great practical importance of these
new processes, and the general recognition of their metallurgical and
commercial value, which has already been accorded to them. It is my
present purpose to review broadly the changes developed by them in
the metallurgy of lead, in which connection it is necessary to refer
briefly to the previous state of the art.
The elimination of the sulphur content of galena has been always the
most troublesome part of the smelting process, being both costly in the
operation and wasteful of silver and lead. Previous to the introduction
of the Huntington-Heberlein process at Pertusola, Italy, it was
effected by a variety of methods. In the treatment of non-argentiferous
galena concentrate, the smelting was done by the roast-reduction method
(roasting in reverberatory furnace and smelting in blast furnace);
the roast-reaction method, applied in reverberatory furnaces; and the
roast-reaction method, applied in Scotch hearths.[34] Precipitation
smelting, simple, had practically gone out of use, although its
reactions enter into the modern blast-furnace practice, as do also
those of the roast-reaction method.
In the treatment of argentiferous lead ores, a combination of the
roast-reduction, roast-reaction and precipitation methods had been
developed. Ores low in lead were still roasted, chiefly in hand-worked
reverberatories (the mechanical furnaces not having proved well adapted
to lead-bearing ores), while the high loss of lead and silver in
sinter-or slag-roasting of rich galenas had caused those processes to
be abandoned, and such ores were charged raw into the blast furnace,
the part of their sulphur which escaped oxidation therein reappearing
in the form of matte. In the roast-reduction smelting of galena alone,
however, there was no way of avoiding the roasting of the whole, or at
least a very large percentage of the ore, and in this roasting the ore
had necessarily to be slagged or sintered in order to eliminate the
sulphur to a satisfactory extent. This is exemplified in the treatment
of the galena concentrate of southeastern Missouri at the present time.
Until the two new Scotch-hearth plants at Alton and Collinsville, Ill.,
were put in operation, the three processes of smelting the southeastern
Missouri galena were about on an equal footing. Their results per ton
of ore containing 65 per cent. lead were approximately as follows[35]:
──────────────────┬──────────────┬────────────
METHOD │ COST │ EXTRACTION
──────────────────┼──────────────┼───────────
Reverberatory │ $6.50-7.00 │ 90-92%
Scotch hearth │ 5.75-6.50 │ 87-88%
Roast-reduction │ 6.00-7.00 │ 90-92%
──────────────────┴──────────────┴───────────
The new works employ the Scotch-hearth process, with bag-houses for
the recovery of the fume, which previously was the weak point of this
method of smelting.[36] This improvement led to a large increase in the
recovery of lead, so that the entire extraction is now approximately 98
per cent. of the content of the ore, while on the other hand the cost
of smelting per ton of ore has been reduced through the increased size
of these plants and the introduction of improved means for handling
ore and material. The practice of these works represents the highest
efficiency yet obtained in this country in the smelting of high-grade
galena concentrate, and probably it cannot be equaled even by the
Huntington-Heberlein and similar processes. The Scotch-hearth and
bag-house process is therefore the one of the older methods of smelting
which will survive.
In the other methods of smelting, a large proportion of the cost is
involved in the roasting of the ore, which amounts in hand-worked
reverberatory furnaces to $2 to $2.50 per ton. Also, the larger
proportion of the loss of metal is suffered in the roasting of the ore,
this loss amounting to from 6 to 8 per cent. of the metal content of
such ore as is roasted. The loss of lead in the combined process of
treatment depends upon the details of the process. The chief advantage
of lime-roasting in the treatment of this class of ore is in the higher
extraction of metal which it affords. This should rise to 98 per cent.
That figure has been, indeed, surpassed in operations on a large scale,
extending over a considerable period.
In the treatment of the argentiferous ores of the West different
conditions enter into the consideration. In the working of those ores,
the present practice is to roast only those which are low in lead,
and charge raw into the blast furnace the rich galenas. The cost of
roasting is about $2 to $2.50 per ton; the cost of smelting is about
$2.50 per ton. On the average about 0.4 ton of ore has to be roasted
for every ton that is smelted. The cost of roasting and smelting is
therefore about $3.50 per ton. In good practice the recovery of silver
is about 98 per cent. and of lead about 95 per cent., reckoned on basis
of fire assays.
In treatment of these ores, the lime-roasting process offers several
advantages. It may be performed at less than the cost of ordinary
roasting.[37] The loss of silver and lead during the roasting is
reduced to insignificant proportion. The sulphide fines which must be
charged raw into the blast furnace are eliminated, inasmuch as they
can be efficiently desulphurized in the lime-roasting pots without
significant loss; all the ore to be smelted in the blast furnace
can be, therefore, delivered to it in lump form, whereby the speed
of the blast furnace is increased and the wind pressure required
is decreased. Finally, the percentage of sulphur in the charge is
reduced, producing a lower matte-fall, or no matte-fall whatever, with
consequent saving in expense of retreatment. In the case of a new
plant, the first cost of construction and the ground-space occupied
are materially reduced. Before discussing more fully the extent and
nature of these savings, it is advisable to point out the differences
among the three processes of lime-roasting that have already come into
practical use.
In the Huntington-Heberlein process, the ore is mixed with suitable
proportions of limestone and silica (or quartzose ore) and is then
partially roasted, say to reduction of the sulphur to one half. The
roasting is done at a comparatively low temperature, and the loss of
metals is consequently small. The roasted ore is dampened and allowed
to cool. It is then charged into a hemispherical cast-iron pot, with
a movable hood which covers the top and conveys off the gases. There
is a perforated grate in the bottom of the pot, on which the ore
rests, and air is introduced through a pipe entering the bottom of the
pot, under the grate. A small quantity of red-hot calcines from the
roasting furnaces is thrown on the grate to start the reaction; a layer
of cold, semi-roasted ore is put upon it, the air blast is turned on
and reaction begins, which manifests itself by the copious evolution
of sulphur fumes. These consist chiefly of sulphur dioxide, but they
contain more or less trioxide, which is evident from the solution of
copperas that trickles from the hoods and iron smoke-pipes, wherein the
moisture condenses. As the reaction progresses, and the heat creeps
up, more ore is introduced, layer by layer, until the pot is full.
Care is taken by the operator to compel the air to pass evenly and
gently through the charge, wherefore he is watchful to close blow-holes
which develop in it. At the end of the operation, which may last from
four to eighteen hours, the ore becomes red-hot at the top. The hood
is then pushed up, and the pot is turned on its trunnions, by means
of a hand-operated wheel and worm-gear, until the charge slides out,
which it does as a solid, semi-fused cake. The pot is then turned back
into position. Its design is such that the air-pipe makes automatic
connection, a flanged pipe cast with the pot settling upon a similarly
flanged pipe communicating with the main, a suitable gasket serving
to make a tight joint. The pots are set at an elevation of about 12
ft. above the ground, so that when the charge slides out the drop will
break it up to some extent, and it is moreover caused to fall on a
wedge, or similar contrivance, to assist the breakage. After cooling it
is further broken up to furnace size by wedging and sledging; the lumps
are forked out, and the fines screened and returned to a subsequent
charge for completion of their desulphurization.
The Savelsberg process differs from the Huntington-Heberlein in respect
to the preliminary roasting, which in the Savelsberg process is
omitted, the raw ore, mixed with limestone and silica, being charged
directly into the converter. The Savelsberg converter is supported on
a truck, instead of being fixed in position, but otherwise its design
and management are quite similar to those of the Huntington-Heberlein
converter. In neither case are there any patents on the converters.
The patents are on the processes. In view of the litigation that
has already been commenced between their respective owners, it is
interesting to examine the claims.
The Huntington-Heberlein patent (U. S. 600,347, issued March 8, 1898,
applied for Dec. 9, 1896) has the following claims:
1. The herein-described method of oxidizing sulphide ores of lead
preparatory to reduction to metal, which consists in mixing with the
ore to be treated an oxide of an alkaline-earth metal, such as calcium
oxide, subjecting the mixture to heat in the presence of air, then
reducing the temperature and finally passing air through the mass
to complete the oxidation of the lead, substantially as and for the
purpose set forth.
2. The herein-described method of oxidizing sulphide ores of lead
preparatory to reduction to metal, which consists in mixing calcium
oxide or other oxide of an alkaline-earth metal with the ore to be
treated, subjecting the mixture in the presence of air to a bright-red
heat (about 700 deg. C.), then cooling down the mixture to a dull-red
heat (about 500 deg. C.), and finally forcing air through the mass
until the lead ore, reduced to an oxide, fuses, substantially as set
forth.
3. The herein-described method of oxidizing lead sulphide in the
preparation of the same for reduction to metal, which consists in
subjecting the sulphide to a high temperature in the presence of an
oxide of an alkaline-earth metal, such as calcium oxide, and oxygen,
and then lowering the temperature substantially as set forth.
Adolf Savelsberg, in U. S. patent 755,598 (issued March 22, 1904,
applied for Dec. 18, 1903) claims:
1. The herein-described process of desulphurizing lead ores, which
consists in mixing raw ore with limestone and then subjecting the
mixture to the simultaneous application of heat and a current of air in
sufficient proportions to substantially complete the desulphurization
in one operation, substantially as described.
2. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with limestone, then
moistening the mixture, then filling it without previous roasting into
a chamber, then heating it and treating it by a current of air, as and
for the purpose described.
3. The herein-described process of desulphurizing lead ores, which
consists in mixing raw ores with limestone, then filling the mixture
into a chamber, then subjecting the mixture to the simultaneous
application of heat and a current of air in sufficient proportions
to substantially complete the desulphurization in one operation, the
mixture being introduced into the chamber in partial charges introduced
successively at intervals during the process, substantially as
described.
4. The herein-described process of desulphurizing lead ores, then
moistening the mixture, then filling it without previous roasting into
a chamber, then heating it and treating it by a current of air, the
mixture being introduced into the chamber in partial charges introduced
successively at intervals during the process, as and for the purpose
described.
5. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with sufficient limestone to
keep the temperature of the mixture below the melting-point of the ore,
then filling the mixture into a chamber, then heating said mixture and
treating it with a current of air, as and for the purpose described.
6. The herein-described process of desulphurizing lead ores, which
process consists in first mixing the ores with sufficient limestone to
mechanically separate the particles of galena sufficiently to prevent
fusion, and to keep the temperature below the melting-point of the ore
by the liberation of carbon dioxide, then filling the mixture into a
chamber, then heating said mixture and treating it with a current of
air, as and for the purpose described.
The Carmichael-Bradford process differs from the Savelsberg by the
treatment of the raw ore mixed with gypsum instead of limestone,
and differs from the Huntington-Heberlein both in respect to the
use of gypsum and the omission of the preliminary roasting. The
Carmichael-Bradford process has not been threatened with litigation,
so far as I am aware. The claims of its original patent read as
follows[38]:
1. The process of treating mixed sulphide ores, which consists in
mixing with said ores a sulphur compound of a metal of the alkaline
earths, starting the reaction by heating the same, thereby oxidizing
the sulphide and reducing the sulphur compound of the alkali metal,
passing a current of air to oxidize the reduced sulphide compound of
the metal of the alkalies preparatory to acting upon a new charge of
sulphide ores, substantially as and for the purpose set forth.
2. The process of treating mixed sulphide ores, which consists in
mixing calcium sulphate with said ores, starting the reaction by
means of heat, thereby oxidizing the sulphide ores, liberating
sulphurous-acid gas and converting the calcium sulphate into calcium
sulphide and oxidizing the calcium sulphide to sulphate preparatory to
treating a fresh charge of sulphide ores, substantially as and for the
purpose set forth.
The process described by W. S. Bayston, of Melbourne (Australian patent
No. 2862), appears to be identical with that of Savelsberg.
Irrespective of the validity of the Savelsberg and Carmichael-Bradford
patents, and without attempting to minimize the ingenuity of their
inventors and the importance of their discoveries, it must be conceded
that the merit for the invention and introduction of lime-roasting of
galena belongs to Thomas Huntington and Ferdinand Heberlein. The former
is an American, and this is the only claim that the United States can
make to a share in this great improvement in the metallurgy of lead. It
is to be regretted, moreover, that of all the important lead-smelting
countries in the world, America has been the most backward in adopting
it.
The details of the three processes and the general results accomplished
by them have been rather fully described in a series of articles
recently published in the _Engineering and Mining Journal_. There
has been, however, comparatively little discussion as to costs; and
unfortunately the data available for analysis are extremely scanty, due
to the secrecy with which the Huntington-Heberlein process, the most
extensively exploited of the three, has been veiled. Nevertheless, I
may attempt an approximate estimation of the various details, taking
the Huntington-Heberlein process as the basis.
The ore, limestone and silica are crushed to pass a four-mesh screen.
This is about the size to which it would be necessary to crush as
preliminary to roasting in the ordinary way, wherefore the only
difference in cost is the charge for crushing the limestone and silica,
which in the aggregate may amount to one-sixth of the weight of the raw
sulphide and may consequently add 2 to 2.5c. to the cost of treating
a ton of ore. The mixing of ore and fluxes may be costly or cheap,
according to the way of doing it. If done in a rational way it ought
not to cost more than 10c. per ton of ore, and may come to less. The
delivery of the ore from the mixing-house to the roasting furnaces
ought to be done entirely by mechanical means, at insignificant cost.
The Heberlein roasting furnace, which is used in connection with the
H.-H. process, is simply an improvement on the old Brunton calciner—a
circular furnace, with revolving hearth. The construction of this
furnace, according to American designs, is excellent. The hearth is
26 ft. in diameter; it is revolved at slow speed and requires about
1.5 h.p. A flange at the periphery of the hearth dips into sand in an
annular trough, thus shutting off air from the combustion chamber,
except through the ports designed for its admittance. The mechanical
construction of the furnace is workmanlike, and the mechanism under the
hearth is easy of access and comfortably attended to.
A 26 ft. furnace roasts about 80,000 lb. of charge per 24 hours. In
dealing with an ore containing 20 to 22 per cent. of sulphur, the
latter is reduced to about 10 to 11 per cent., the consumption of
coal being about 22.5 per cent. of the weight of the charge. The
hearth efficiency is about 150 lb. per sq. ft., which in comparison
with ordinary roasting is high. The coal consumption, however, is not
correspondingly low. Two furnaces can be managed by one man per 8 hour
shift. On the basis of 80 tons of charge ore per 24 hours, the cost of
roasting should be approximately as follows:
Labor—3 men at $2.50 $ 7.50
Coal—18 tons at $2 36.00
Power 3.35
Repairs 3.35
——————
Total $50.20 = 63c. per ton.
In the above estimate repairs have been reckoned at the same figure as
is experienced with Brückner cylinders, and the cost of power has been
allowed for with fair liberality. The estimated cost of 63c. per ton
is comparable with the $1.10 to $1.45 per ton, which is the result of
roasting in Brückner cylinders in Colorado, reducing the ore to 4.5-6
per cent. sulphur.
The Heberlein furnace is built up to considerable elevation above
the ground level, externally somewhat resembling the Pearce turret
furnace. This serves two purposes: (1) it affords ample room under the
hearth for attention to the driving mechanism; and (2) it enables the
ore to be discharged by gravity into suitable hoppers, without the
construction of subterranean gangways. The ore discharges continuously
from the furnace, at dull-red heat, into a brick bin, wherein it is
cooled by a water-spray. Periodically a little ore is diverted into a
side bin, in which it is kept hot for starting a subsequent charge in
the converter.
The cooled ore is conveyed from the receiving bins at the roasting
furnaces to hopper-bins above the converters. If the tramming be done
by hand the cost, with labor at 25c. per hour, may be approximately
12.5c. per ton of ore, but this should be capable of considerable
reduction by mechanical conveyance.
The converters are hemispherical pots of cast iron, 9 ft. in
diameter at the top, and about 4 ft. in depth. They are provided
with a circular, cast-iron grate, which is ¾ in. thick and 6 ft. in
diameter and is set and secured horizontally in the pot. This grate
is perforated with holes ¾ in. in diameter, 2 in. apart, center to
center, and is similar to the Wetherill grate employed in zinc oxide
manufacture. The pot itself is about 2½ in. thick at the bottom,
thinning to about 1½ in. at the rim. It is supported on trunnions and
is geared for convenient turning by hand. The blast pipe which enters
the pot at the bottom is 6 in. in diameter.
Two roasting furnaces and six converters are rated nominally as a 90
ton plant. This rating is, however, considerably in excess of the
actual capacity, at least on certain ores. The time required for
desulphurization in the converter apparently depends a good deal upon
the character of the ore. The six converters may be arranged in a
single row, or in two rows of three in each. They are set so that the
rim of the pot, when upright, is about 12 ft. above the ground level.
A platform gives access to the pots. One man per shift can attend to
two pots. His work consists in charging them, which is done by gravity,
spreading out the charge evenly in the pot, closing any blow-holes
which may develop, and at the end of the operation raising the hood
(which covers the pot during the operation) and dumping the pot. The
work is easy. The conditions under which it is done are comfortable,
both as to temperature and atmosphere. Reports have shown a great
reduction in liability to lead-poisoning in the works where the H.-H.
process has been introduced.
A new charge is started by kindling a small wood or coal fire on the
grate, then throwing in a few shovelfuls of hot calcines, and finally
dropping in the regular charge of damp ore (plus the fluxes previously
referred to). The charge is introduced in stages, successive layers
being dropped in and spread out as the heat rises. At the beginning
the blast is very low—about 2 oz. It is increased as the hight of the
ore in the pot rises, finally attaining about 16 oz. The operation
goes on quietly, the smoke rising from the surface evenly and gently,
precisely as in a well-running blast furnace. While the charge is still
black on top, the hand can be held with perfect comfort, inside of
the hood, immediately over the ore. This explains, of course, why the
volatilization of silver and lead is insignificant. There is, moreover,
little or no loss of ore as dust, because the ore is introduced damp,
and the passage of the air through it is at low velocity. In the
interior of the charge, however, there is high temperature (evidently
much higher than has been stated in some descriptions), as will be
shown further on. The conditions in this respect appear to be analogous
to those of the blast furnace, which, though smelting at a temperature
of say 1200 deg. C. at the level of the tuyeres, suffers only a slight
loss of silver and lead by volatilization.
At the end of the operation in the H.-H. pot, the charge is dull red
at the top, with blow-holes, around which the ore is bright red.
Imperfectly worked charges show masses of well-fused ore surrounded
by masses of only partially altered ore, a condition which may be
ascribed to the irregular penetration of air through the charge,
affording good evidence of the important part which air plays in the
process. A properly worked charge is tipped out of the pot as a solid
cake, which in falling to the ground breaks into a few large pieces.
As they break, it appears that the interior of the charge is bright
red all through, and there is a little molten slag which runs out of
cavities, presumably spots where the chemical action has been most
intense. When cold, the thoroughly desulphurized material has the
appearance of slag-roasted galena. Prills of metallic lead are visible
in it, indicating reaction between lead sulphide and lead sulphate.
The columns of the structure supporting the pots should be of steel,
since fragments of the red-hot ore dumped on the ground are likely to
fall against them. To hasten the cooling of the ore, water is sometimes
played on it from a hose. This is bad, since some is likely to splash
into the still inverted pot, leading to cracks. The cracked pots at
certain works appear to be due chiefly to this cause, in the absence of
which the pots ought to last a long time, inasmuch as the conditions
to which they are subjected during the blowing process are not at all
severe. When the ore is sufficiently cold it is further broken up,
first by driving in wedges, and finally by sledging down to pieces
of orange size, or what is suitable for the blast furnace. These are
forked out, leaving the fine ore, which comes largely from the top of
the charge and is therefore only partially desulphurized. The fines
are, therefore, re-treated with a subsequent charge. The quantity is
not excessive; it may amount to 7 or 8 per cent. of the charge.
The breaking up of the desulphurized ore is one of the problems of the
process, the necessity being the reduction of several large pieces
of fused, or semi-fused, material weighing two or three tons each.
When done by hand only, as is usually (perhaps always) the practice,
the operation is rather expensive. It would appear, however, to
be not a difficult matter to devise some mechanical aids for this
process—perhaps to make it entirely mechanical. When done by hand, a
6-pot plant requires 6 men per shift sledging and forking. With 8-hour
shifts, this is 18 men for the breaking of about 60 tons of material,
which is about 3⅓ tons per man per 8 hours. With labor at 25c. per
hour, the cost of breaking the fused material comes to 60c. per ton. It
may be remarked, for comparison, that in breaking ore as it ordinarily
comes, coarse and fine together, a good workman would normally be
expected to break 5 to 5.5 tons in a shift of 8 hours.
The ordinary charge for the standard converter is about 8 tons (16,000
lb.) of an ore weighing 166 lb. per cu. ft. With a heavier ore, like
a high-grade galena, the charge would weigh proportionately more. The
time of working off a charge is decidedly variable. Accounts of the
operation of the process in Australia tell of charge-workings in 3
to 5 hours, but this does not correspond with the results reported
elsewhere, which specify times of 12 to 18 hours. Assuming an average
of 16 hours, which was the record of one plant, six converters would
have capacity for about 72 tons of charge per 24 hours, or about 58
tons of ore, the ratio of ore to flux being 4:1. The loss in weight
of the charge corresponds substantially to the replacement of sulphur
by oxygen, and the expulsion of carbon dioxide. The finished charge
contains on the average from 3 to 5 per cent. sulphur. This is
about the same as the result achieved in good practice in roasting
lead-bearing ores in hand-worked reverberatory furnaces, but curiously
the H.-H. product, in some cases at least, does not yield any matte,
to speak of, in the blast furnace; the product delivered to the latter
being evidently in such condition that the remaining sulphur is almost
completely burned off in the blast furnace. This is an important saving
effected by the process. In calculating the value of an ore, sulphur
is commonly debited at the rate of 25c. per unit, which represents
approximately the cost of handling and reworking the matte resulting
from it. The practically complete elimination of matte-fall rendered
possible by the H.-H. process may not be, however, an unmixed blessing.
There may be, for example, a small formation of lead sulphide which
causes trouble in the crucible and lead-well, and results in furnace
difficulties and the presentation of a vexatious between-product.
It may now be attempted to summarize the cost of the converting
process. Assuming the case of an ore assaying lead, 50 per cent.; iron,
15; sulphur, 22; silica, 8, and alumina, etc., 5, let it be supposed
that it is to be fluxed with pure limestone and pure quartz, with the
aim to make a slag containing silica, 30; ferrous oxide, 40; and lime,
20 per cent. A ton of ore will make, in round numbers, 1000 lb. of
slag, and will require 344 lb. of limestone and 130 lb. of quartz,
or we may say roughly one ton of flux must be added to four tons of
ore, wherefore the ore will constitute 80 per cent. of the charge. In
reducing the charge to 3 per cent. sulphur it will lose ultimately
through expulsion of sulphur and carbon dioxide (of the limestone)
about 20 per cent. in weight, wherefore the quantity of material to
be smelted in the blast furnace will be practically equivalent to
the raw sulphide ore in the charge for the roasting furnaces; but in
the roasting furnace the charge is likely to gain weight, because of
the formation of sulphates. Taking the charge, which I have assumed
above, and reckoning that as it comes from the roasting furnace it
will contain 10 per cent. sulphur, all in the form of sulphate, either
of lead or of lime, and that the iron be entirely converted to ferric
oxide, in spite of the expulsion of the carbon dioxide of the limestone
and the combustion of a portion of the sulphur of the ore as sulphur
dioxide, the charge will gain in weight in the ratio of 1:1.19. This,
however, is too high, inasmuch as a portion of the sulphur will remain
as sulphide while a portion of the iron may be as ferrous oxide. The
actual gain in weight will consequently be probably not more than
one-tenth. The following theoretical calculation will illustrate the
changes:
─────────────────────┬──────────────────────┬─────────────────────────
RAW CHARGE │ SEMI-ROASTED CHARGE │ FINISHED CHARGE
─────────────────────┼──────────────────────┼─────────────────────────
{1000 lb. Pb │ {1154 lb. PbO │ { 1154 lb. PbO
{ 300 lb. Fe │ { 428 lb. Fe₂O₃ │ { 428 lb. Fe₂O₃(?)
Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂ │ Ore { 160 lb. SiO₂
{ 100 lb. Al₂O₃,│ { 100 lb. Al₂O₃, │ { 100 lb. Al₂O₃,
etc. │ etc. │ etc.
{ 440 lb. S │ { 300 lb. S │ { 68 lb. S
│ │
{ 130 lb. SiO₂ │ { 130 lb. SiO₂ │ { 130 lb. SiO₂
Flux { 344 lb. CaCO₃ │ Flux { 193 lb. CaO │ Flux { 193 lb. CaO
│ 450 lb. O │
———— │ ———— │ ————
2474 lb. │ 2915 lb. │ 2233 lb.
│ │
│ 10% S. │ 3% S.
─────────────────────┴──────────────────────┴─────────────────────────
Ratios:
2474:2915 :: 1:1.18.
2915:2233 :: 1:0.76⅔.
2474:2233 :: 1:0.90.
It may be assumed that for every ton of charge (containing about 80 per
cent. of ore) there will be 1.1 ton of material to go to the converter,
and that the product of the latter will be 0.9 of the weight of the
original charge of raw material.
Each converter requires 400 cu. ft. of air per minute. The blast
pressure is variable, as different pots are always at different stages
of the process, but assuming the maximum of 16 oz. pressure, with a
blast main of sufficient diameter (at least 15 in.) and the blower
reasonably near the battery of pots, the total requirement is 21 h.p.
The cost of converting will be approximately as follows:
Labor, 3 foremen at $3.20 $ 9.60
“ 9 men at $2.50 22.50
Power, 21 h.p. at 30c 6.30
Supplies, repairs and renewals 5.00
——————
Total $43.40 = 60c. per ton of charge.
The cost of converting is, of course, reduced directly as the time is
reduced. The above estimate is based on unfavorable conditions as to
time required for working a charge.
The total cost of treatment, from the initial stage to the delivery of
the desulphurized ore to the blast furnaces, will be, per 2000 lb. of
charge, approximately as follows:
Crushing 1.0 ton at 10c $0.10
Mixing 1.0 ton at 10c .10
Roasting 1.0 ton at 63c .63
Delivering 1.1 ton to converters at 12c .13
Converting 1.1 ton at 60c .66
Breaking 0.9 ton at 60c .54
——-——-
Total $2.16
The cost per ton of ore will be 2.16 ÷ 0.80 = $2.70. Making allowance
for the crushing of the ore, which is not ordinarily included in the
cost of roasting, and possibly some overestimates, it appears that the
cost of desulphurization by this method, under the conditions assumed
in this paper, is rather higher than in good practice with ordinary
hand-worked furnaces, but it is evident that the cost can be reduced to
approximately the same figure by introduction of improvements, as for
example in breaking the desulphurized ore, and by shortening the time
of converting, which is possible in the case of favorable ores. The
chief advantage must be, however, in the further stage of the smelting.
As to this, there is the evidence that the Broken Hill Proprietary
Company was able to smelt the same quantity of ore in seven furnaces,
after the introduction of the Huntington-Heberlein process, that
formerly required thirteen. A similar experience is reported at
Friedrichshütte, Silesia.
This increase in the capacity of the blast furnace is due to three
things: (1) In delivering to the furnace a charge containing a reduced
percentage of fine ore, the speed of the furnace is increased, i.e.,
more tons of ore can be smelted per square foot of hearth area. (2)
There is less roasted matte to go into the charge. (3) Under some
conditions the percentage of lead in the charge can be increased,
reducing the quantity of gangue that must be fluxed.
It is difficult to generalize the economy that is effected in the
blast-furnace process, since this must necessarily vary within wide
limits because of the difference in conditions. An increase of 60 to
100 per cent. in blast-furnace capacity does not imply a corresponding
reduction in the cost of smelting. The fuel consumption per ton of ore
remains the same. There is a saving in the power requirements, because
the smelting can be done with a lower blast pressure; also, a saving
in the cost of reworking matte. There will, moreover, be a saving in
other labor, in so far as portions thereof are not already performed
at the minimum cost per ton. The net result under American conditions
of silver-lead smelting can only be determined closely by extensive
operations. That there will be an important saving, however, there is
no doubt.
The cost of smelting a ton of charge at Denver and Pueblo, exclusive
of roasting and general expense, is about $2.50, of which about $0.84
is for coke and $1.66 for labor, power and supplies. General expense
amounts to about $0.16 additional. If it should prove possible to
smelt in a given plant 50 per cent. more ore than at present without
increase in the total expense, except for coke, the saving per ton of
charge would be 70c. That is not to be expected, but the half of it
would be a satisfactory improvement. With respect to sulphur in the
charge, the cost is commonly reckoned at 25c. per unit. As compared
with a charge containing 2 per cent. of sulphur there would be a saving
rising toward 50c. per ton as the maximum. It is reasonable to reckon,
therefore, a possible saving of 75c. per ton of charge in silver-lead
smelting, no saving in the cost of roasting, and an increase of about
3 per cent. in the extraction of lead, and perhaps 1 per cent. in the
extraction of silver, as the net results of the application of the
Huntington-Heberlein process in American silver-lead smelting.
On a charge averaging 12 per cent. lead and 33 oz. silver per ton,
an increase of 3 per cent. in the extraction of lead and 1 per
cent. in the extraction of silver would correspond to 25c. and 35c.
respectively, reckoning lead at 3.5c. per lb., and silver at 60c. per
oz. In this, however, it is assumed that all lead-bearing ores will
be desulphurized by this process, which practically will hardly be
the case. A good deal of pyrites, containing only a little lead, will
doubtless continue to be roasted in Brückner cylinders, and other
mechanical furnaces, which are better adapted to the purpose than are
the lime-roasting pots. Moreover, a certain proportion of high-grade
lead ore, which is now smelted raw, will be desulphurized outside of
the furnace, at additional expense. It is comparatively simple to
estimate the probable benefit of the Huntington-Heberlein process in
the case of smelting works which treat principally a single class of
ore, but in such works as those in Colorado and Utah, which treat a
wide variety of ores, we must anticipate a combination process, and
await results of experience to determine just how it will work out.
It should be remarked, moreover, that my estimates do not take into
account the royalty on the process, which is an actual debit, whether
it be paid on a tonnage basis or be computed in the form of a lump sum
for the license to its use.
However, in view of the immense tonnage of ore smelted annually for
the extraction of silver and lead, it is evident that the invention of
lime-roasting by Huntington and Heberlein was an improvement of the
first order in the metallurgy of lead.
In the case of non-argentiferous galena, containing 65 per cent. of
lead (as in southeastern Missouri), comparison may be made with the
slag-roasting and blast-furnace smelting of the ore. Here, no saving
in cost of roasting may be reckoned and no gain in the speed of the
blast furnaces is to be anticipated. The only savings will be in
the increase in the extraction of lead from 92 to 98 per cent., and
the elimination of matte-roasting, which latter may be reckoned as
amounting to 50c. per ton of ore. The extent of the advantage over
the older method is so clearly apparent that it need not be computed
any further. In comparison with the Scotch-hearth bag-house method of
smelting, however, the advantage, if any, is not so certain. That
method already saves 98 per cent. of the lead, and on the whole is
probably as cheap in operation as the Huntington-Heberlein could be
under the same conditions. The Huntington-Heberlein method has replaced
the old roast-reaction method at Tarnowitz, Silesia, but the American
Scotch-hearth method as practised near St. Louis is likely to survive.
A more serious competitor will be, however, the Savelsberg process,
which appears to do all that the Huntington-Heberlein process does,
without the preliminary roasting. Indeed, if the latter be omitted
(together with its estimated expense of 63c. per ton of charge, or
79c. per ton of ore), all that has been said in this paper as to the
Huntington-Heberlein process may be construed as applying to the
Savelsberg. The charge is prepared in the same way, the method of
operating the converters is the same, and the results of the reactions
in the converters are the same. The litigation which is pending between
the two interests, Messrs. Huntington and Heberlein claiming that
Savelsberg infringes their patents, will be, however, a deterrent to
the extension of the Savelsberg process until that matter be settled.
The Carmichael-Bradford process may be dismissed with a few words. It
is similar to the Savelsberg, except that gypsum is used instead of
limestone. It is somewhat more expensive because the gypsum has to be
ground and calcined. The process works efficiently at Broken Hill,
but it can hardly be of general application, because gypsum is likely
to be too expensive, except in a few favored localities. The ability
to utilize the converter gases for the manufacture of sulphuric acid
will cut no great figure, save in exceptional cases, as at Broken
Hill, and anyway the gases of the other processes can be utilized for
the same purpose, which is in fact being done in connection with the
Huntington-Heberlein process in Silesia.
The cost of desulphurizing a ton of galena concentrate by the
Carmichael-Bradford process is estimated by the company controlling
the patents as follows, labor being reckoned at $1.80 per eight hours,
gypsum at $2.40 per 2240 lb., and coal at $8.40 per 2240 lb.:
0.25 ton of gypsum $0.60
Dehydrating and granulating gypsum .48
Drying mixture of ore and gypsum .12
Converting 0.24
Spalling sintered material .12
0.01 ton coal .08
——-——-
Total $1.64
The value of the lime in the sintered product is credited at 12c.,
making the net cost $1.52 per 2240 lb. of ore.
The cost allowed for converting may be explained by the more rapid
action that appears to be attained with the ores of Broken Hill than
with some ores that are treated in North America, but the low figure
estimated for spalling the sintered material appears to be highly
doubtful.
The theory of the lime-roasting processes is not yet well established.
It is recognized that the explanation offered by Huntington and
Heberlein in their original patent specification is erroneous. There is
no good evidence in their process, or any other, of the formation of
the higher oxide of lime, which they suggest.
At the present time there are two views. In one, formulated most
explicitly by Professor Borchers, there is formed in this process a
plumbate of calcium, which is an active oxidizing agent. A formation of
this substance was also described by Carmichael in his original patent,
but he considered it to be the final product, not the active oxidizing
agent.
In the other view, the lime, or limestone, serves merely as a diluent
of the charge, enabling the air to obtain access to the particles of
galena, without liquefaction of the latter. The oxidation of the lead
sulphide is therefore effected chiefly by the air, and the process
is analogous to what takes place in the bessemer converter or in the
Germot process of smelting, or perhaps more closely to what might
happen in an ordinary roasting furnace, provided with a porous hearth,
through which the air supply would be introduced. Roasting furnaces of
that design have been proposed, and in fact such a construction is now
being tested for blende roasting in Kansas.
Up to the present time, the evidence is surely too incomplete to enable
a definite conclusion to be reached. Some facts may, however, be stated.
There is clearly reaction to a certain extent between lead sulphide
and lead sulphate, as in the reverberatory smelting furnace, because
prills of metallic lead are to be observed in the lime-roasted charge.
There is a formation of sulphuric acid in the lime-roasting, upon the
oxidizing effect of which Savelsberg lays considerable stress, since
its action is to be observed on the iron work in which it condenses.
Calcium sulphate, which is present in all of the processes, being
specifically added in the Carmichael-Bradford, evidently plays an
important chemical part, because not only is the sulphur trioxide
expelled from the artificial gypsum, but also it is to a certain
extent expelled from the natural gypsum, which is added in the
Carmichael-Bradford process; in other words, more sulphur is given off
by the charge than is contained by the metallic sulphides alone.
Further evidence that lime does indeed play a chemical part in the
reaction is presented by the phenomena of lime-roasting in clay dishes
in the assay muffle, wherein the air is certainly not blown through the
charge, which is simply exposed to superficial oxidation as in ordinary
roasting.
The desulphurized charge dropped from the pot is certainly at much
below the temperature of fusion, even in the interior, but we have no
evidence of the precise temperature condition during the process itself.
Pyrite and even zinc blende in the ore are completely oxidized. This,
at least, indicates intense atmospheric action.
The papers by Borchers,[39] Doeltz,[40] Guillemain,[41] and
Hutchings[42] may profitably be studied in connection with the
reactions involved in lime-roasting. The conclusion will be, however,
that their precise nature has not yet been determined. In view of the
great interest that has been awakened by this new departure in the
metallurgy of lead, it is to be expected that much experimental work
will be devoted to it, which will throw light upon its principles, and
possibly develop it from a mere process of desulphurization into one
which will yield a final product in a single operation.