Charles George Warnford Lock.

Economic mining: a practical handbook for the miner, the metallurgist and ... online

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oantle, so that the interior portion, composed of firebricks, can, when
nimt out, be easily removed without disturbing the superstructure,
rhe section of the furnace widens upwards towards the feed doors,
rhich arrangement is exceedingly advantageous, in view of the fact
hat the charge becomes compacted as it descends towards the smelt-
Dg zone. The gases as they ascend to the upper parts of the stack
lave an opportunity to expand, thus diminishing their velocity, and
or this reason the amount of flue dust is very considerably lessened.
The Baschette is particularly applicable to smelting both lead and
opper ores, and as originally constructed, had two working fronts,
ind was proportionately very much longer than it was wide. It was
bund that this furnace put through in 24 hours 40-60 per cent, more
»re than the round furnaces. The upper portion was, like the Pilz,
npported upon an iron mantle resting upon iron columns. It has
brmed the oasis of modem improvements. To prevent the loss of
ime, labour, and temperature occasioned by the frequent burning
hrough of the lower portion of the furnace, the brickwork in that
Murt has been replaced by a so-called "water jacket," an annular
ylinder of iron about 3 ft. high, which is kept cool by a constant
itream of cold water running through it.

The essential features of a typical lead smelting furnace of t.o-day
ire shown in Fig. 155. The water jacket H is either cast in one piece
K constructed of j^in. boiler plate. The position and number of the
nijers b is a matter of importance. The usual number is 5, which
)ieroe at equal distances the lower third of the water jacket and con-
verge towiu>ds the centre of the furnace. By reducing the number of
nyers, or placing them farther from the breast, the water would cool
iie interior of the jacket to such a degree as to interfere with the
regular descent of the charge. The water enters through an inlet
pipe at the bottom of the jacket, supplied with a valve to regulate
the supply, and leaves on the opposite side near the upper edge of the
jacket. The arrangement shown in dotted lines at c can be recom-
tnended, as the workman may readily estimate the quantity and
temperature of the water as it falls firom the outlet pipes into the
funnel c communicating with the drain, and reeulate the cold water
mpply accordingly. The upper part of the fomace is frequently
encased in sheet iron, strongly riveted together, to strengthen it and
* See J. IMby, ^ Argentiferoas Lead liines of Spain," in lnduUrie$,

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S38 ECONOMIC MINING.

to prevent the escape of gases. A sheet-iron hood d is placed 0¥«
the fore hearth e, and carries off lead fames escaping from the breitft
and thus preyents them from injurionsly affecting the health of tin
charger ahove. When necessary, this hood may he pulled np, hj
means of a chain and pnlley /, so as not to interfere with the woiIk
The charging is done at the top, which is preferable to chajrging firon
the side or rear, as less atmospneric air enters the flue. The bottoo
of the fomaoe is made of brasque, and hollowed in the usual w^y t
form a cavity for the collection of the melted lead. At ^ ib th
automatic or siphon tap or lead well invented by Reyes and Arenti



Fio. 155.— MoDKBH Blast Fubnagi.

It should be so constructed that a bar inserted from the outside wd
readily pass to the bottom of tiie furnace, an angle of 35^-45^ beiii|
best. Some further points which Reyes * thinks would be improve
ments are :— (a) that the interior walls of the furnace, instead of pro
ceeding directly upward, should expand in the form of a letter V, '
lower portion being towards the hearth ; (b) a forward rake of
tuyers towards the hearth or slag discharge of the furnace w<
drive the slag as formed towards its natiural exit, the slag d(
(c) provision of two spouts, one 2-2^ in. higher than the other
the discharge of the slag, the lower being intended for discharge
matte or speiss, if either or both should be formed. His ideal fm
would have a hearth area of 8J^ x 3^ ft., with 11 tuyers, 5 on
side and 1 in the back wall, nozzles 4 in. diam., and inserted 10

* W. a Keyes, ** Notes on Western Lead Smelting,'' Rep. Stato Mlnermlodl
\if., 1888. ^^



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METALLIFEROUS MINERALS. 539

ibove upper edge of hearth plate. Furnace to be 13 ft. high from
»ntre of tnyers to feed door; and cracible 26 in. deep below edge of
iiearth plate. Blowers should be in engine room under supervision
>f engineer* and drawing air from outside through a brick conduit,
is to wood charcoal, it is necessary to study the characteristics of
Ufierent kinds before adoption: thus Eeyes found the *' mountain
nahogany^ charcoal crush to impalpable powder after combustion
md actually put out the fire. He also instances an experiment in
naking aluminous slag when silidous flux was wanting, employing
day slate for the purpose with success.

Cironlar furnaces are giving way to oblong,* because their smelting
sapocity is limited by the diameter at the tuyers, which practice has
ixed at 42 in. maximum, commonly 36 in. The oblong furnace at
uyor level is 86-120 in. long and 30-42 in. wide, the two figures for
iridth representing the views of opposite schools, low pressure (|-1 in.
neroory) and high pressure (2-2j^ in.) blasts. The strength of blast
letermines the height between tuyers and feed floor (12-18 ft.). A
urnaoe 100 x 33 in. at the tuyers, with 5 tuyers (f in.) on each side.
Old 12 ft. active height, at 1^ in. pressure, will smelt 60 tons of
nedioin charge. It is an improvement to have the lead well enclosed
Tj the cruci me castings and not bolted to them ; and cast-iron tuyer-
)oxe8 bolted to the jackets are superior to sheet-iron tuyer-pipes.
lofinan believes in thej possible " replacing of the water jackets
rhoUy or in part by a suitable refractory material," as the cooling
vater consumes a large amount of heat, e.g. a furnace 36 by 92 in. at
lie tuyers, requires per minute 11 gal. water, the temperature of
irhic^ becomes raised from say 60^ to 160° F. Tlie desired refractory
naterial may prove to be coke-brick. The fuels used range &om
sharcoal and coke to bituminous and anthracite coals. In America
hib long double hearth reverberatory 14-16 ft wide is preferred, the
ve from the roasting hearth falling vertically 22-24 m. on to the
Jagging hearth. As low as 10 per cent, ores are worked. A furnace
K) X 14 ft will slag-roast 2-3 tons of charge per 24 hours. Becovery
reaches 94 per cent of the lead in poor ores, and 95 per oent. of the
olver in almost alL Cost is approximately 8«. a ton for calcining, and
^ a ton for smelting, assuming a furnace to take 48 tons ore per
)A hours, using 33 per oent. flux (16 tons at 12«.) and 18 per cent rael
coke at 48«., and 3 cords wood at 24«.), and labour, 4&c., amounting
X)25I.

At Broken Hill, the charge consists of h\\ per cent lead ore,
\\ iron ore, 47 silicious iron and kaolin, 18 coke, and 32 limestone.
Hie limestone, broken to 4 in., and guaranteed 94 per cent, lime
sarbonate, is delivered at 18«. a ton. English coke is used, costing
tieavhy and suffering much from handling. Total cost of smelting is
kbout 32«. a ton, including 20«. for fael, 7«. 6d. for fluxes, and 7«.
for labour ; and the grand total cost of mining, reduction, and realisa-
tion is about 32. 14«. a ton.

The usual plan for preparing the smelting mixture is to lay down
I large number of tons together with the proper fluxes, and to weigh

* H. O. Hofnum, ** Treatment of Argon tiforous Load Ores," Mineral Industry,
1321; 11427.



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S40 ECONOMIC MINING.

out the charges at the feed door. Where the ores are of a diffiBi«ii1
sifee and different character, it may be better to have the feeder wmgli
out directly the charge, or half charge of ore and flnx^, just befoit
patting them into the furnace. This requires more labour, but h
often sufficiently advantageous to warrant the increased expense.

All ores are either acid, basic, or neutral, the last so seldom diai
it may be left out. The principal constituents to be regarded in tb<
matter of preparing admixtures for the formation of a proper slag are
(a) silica, representing the acid; (&) iron, lime, magnesia, and tbe
alkalies. Alumina is sometimes present and acts either as a base oi
as an acid. It is usually considered as equivalent to silica anj
reckoned as such. Magnesia and baryta when present are reduced t(
lime in the ratio of their molecular weights, and entered in the calcu-
lations as calcium oxide: thus, MgO X 1*4, BaO X *4, and, by sonu
smelters, ZnO X *7. The alkalies being usually in small proportioD
may be disregarded or allowed for amongst the bases. In order ti
calculate the constituents of a slag at all, it is necessary to have fol
and complete analyses of the ore and fluxes. Knowing the oompofi
tion of both the ores and fluxes, the head smelter can proceed to tin
combination of a slag which will meet, not only the metallurgioa]
but the economical conditions of his locality. The principal slags an
the following : —

(a) Sub-silicates, where the oxygen of the base is to the oxygei
of the silica, the acid, as 2 to 1 ; chemical formula, 4Ro,Si02.

(&) SinguIoHsilicates, in which the oxygen of the base is to th
oxygen of the acid, the silica, as 1 to 1 ; chemical formula, 2 Bo
SiOj.

(c) Bi-silicates, in which the oxygen of the base is to the oxygei
of the silica, the acid, as 1 to 2 ; chemical formula, BoSiO,.

In practice these slags are mingled both mechanically and ohemi
cally in various proportions, and of such admixtures the following J
are of common occurrence :* —

(a) 1 mono-silicate of iron plus 5 bi-silicates of lime ; ohemica
formula, (2FeO,SiOjj)+6(('aOSi02); percentage— 46 SiO„ 18 FeO
36 GaO. Such a composition is advisable when iron is soaroe anc
silicious ore plentiful.

(b) 1 mono-silicate of iron, 1 bi-silicate of iron, with 2 monoj
silicates of lime; chemical formula, j2FeO,Si02,FeOSiO,}-|.2(2C«C^
SiOa) ; percentage— 35i SiO^, 31^ FeO, 33 Caa

(c) 1 mono-silicate of iron plus 1 mono-silicate of lime ; ohemica]
formula, 2FeO,SiOa-h2CaO,Si02; percentage— 32 SiO„ 38i FeO|
29} CaO.

(d) 1 mono-silicate of iron, 1 bi-silicate of iron, and 1 mono-aUioat^
of lime; chemical formula, {2FeO,Si03,FeOSi02}4-2CaO,SiO,; peri
centage— 35^ SiO,, 42j^ FeO, 22 GaO ; this is, as a rule, the best ^
or the common run of ores.

(e) 3 bi-silicate of iron, 1 bi-silicate of lime, plus 3 mono-silicate
of iron and 1 mono-silicate of lime; chemical formula, 3FeO8i0j
CaOSi02-f3(2FeO,SiOa),2CaO,SiOa; percentage— 37 SiO„ 60 FeO,
13 GaO ; this was first extensively used by A. Eilers, and is partioH

* Bep. State Mineralogist Calif., 1890, p. 829.

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METALLIFEROUS MINERALS. 541

larly to be commended when the ore is of a highly fermginotis nature
fas it requires little quartz and lime), also for use when zinc-bearing
Dres are to be reduced. In all these slags some allowance has to be
made for the alkalies. Slags rich in lime will cany and neutralise a
greater percentage of silica than those of iron, without taking up too
nuch oxide of lead. Basic slags cause overfire or flaming at the
diroat, and have a tendency to corrode the brickwork of the furnace,
ind to cause loss of metal owing to the rapidity and ease of their
ormation. A too acid slag, on the contrary, retards the fusibility of
he furnace mixtures, and hence the furnace runs too slow. The ore,
18 receiTcd, of course contains moisture, and in calculating the charge
t is necessary to make the proper modification, for the reason that
he ore and flux analysis is based upon a steam-dried material It is
iirther to be observed that 31 ' 05 iron requires 34 silica (SiOj) for its
leutralisation, and that 16 sulphur requires 28 metallic iron. Having
he analyses of the ore, flux and fuel, the metallurgist can proceed to
alculate the percentages required either to supply what is lacking in
he ore, or to render harmless an excess of acid or sulphur ; it becomes
imply a question of mathematical proportion. As to the amounts of
^ and silver in the respective charges, the smelter must be
ovemed by circumstances and his practical experience. About
00 oz. silver to the ton of lead appears to be the most desirable pro-
ortion to " cover " the silver and prevent loss.

The conditions laid down by Collins* are somewhat different,
fe insists that : — (a) the slags must never average above 40 per cent.
iOj, and, if fluxiog ores are available, especially those with large
•on excess, it is better to aim at a slag with only 32 per cent. SiOj ;
») the amount of lead in the charge should be such as to yield 10 per
snt. of the weight of the latter as lead bullion, irrespective of the
nail amount contained in the matte ; (e) the bullion snould not run
inch above 300 oz. silver per ton. Special cases often arise in which
becomes necessary to use for a time slags of 42 per cent. SiO,, or to
ork with a bullion-yield of only 8 per cent., or to produce bullion
r upwards of 400 oz. per ton; but such conditions are always
atrimental.

The basis of calculations for slags is that the most fusible of all
on silicates is the mono-silicate (2FeO,Si02), or 70 per cent, ferrous
Lide and 30 silica; and though the most fusible of the calcium
licates is the bisilicate (CaO,Si02), or 63 per cent, silica and 47 lime,
3t the mono-silicate is preferable because the extra lime helps carry
r sulphur as calcium sulphide in the slags. Therefore, all the iron
calculated as 2FeO,Si02, and all the lime as 2CaO,Si02 ; excess of
Ilea being reckoned separately, and neutralised wiUi limestone,
iless argentiferous iron ore can be got.

In valuing an ore for smelting its composition has to be carefully
^nsidered. The presence of over 6 per cent, zinc in sulphuretted
66 and over 10 per cent, in oxidised, entails a reduction of about 2«.
mut for all excess per ton. For barytes beyond 10 per cent., a
\xrwd of 7-8c2. a unit is made. When the ores are calcareous or
ghly charged with iron oxide, a reduction is often made in the
♦ H. F. Collins, op. cit



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542 ECONOMIC MINING.

oliarge for smelting. As regards the lead, some smelten reqnin a
certain minimum, and an extra charge is made for each nnit of lead
below such minimum. It is nsnal to dednct 5 per cent, for lost of
silver in ore, bnt if the ores are dry or silicions, a still farther deduc-
tion is made, often as high as 10 per cent, for 100 oz. ore, and lees for
the higher grades. The usual charge for smelting dry ores in Colondo
is 21. 10«.-3/. and eyen 5/. a ton. Gold beyond ^ os. is paid for at do
per cent. When lead goes less than 5-10 per cent, no payment »
made for it. Excess of silica is charged at 5d. per unit. Copper u
not paid for in lead ores nor lead in copper ores.

Purification. — Lead as delivered by the various smelting prooeeaes
contains generally, in addition to silver, some proportion of antimooy,
arsenic, copper, iron and zinc, which may be in such amount as to in-
terfere witn recovery of the silver. The first step therefore is their
more or less perfect removal, b^ which the lead is " softened '* or
** improved." The operation consists in melting the crude lead (" wcrk
lead" or '* base bullion") and exposing it to the oxidising influence of
the atmosphere until all impurities have been removed as scum, as^
a pure lead-silver alloy remains behind. The apparatus required oos-
sists in a melting pot and an oxidising pan, both made of cast inm.
The oxidising pan is generally 10-12 ft. long, 5-6 ft. wide, and about
10 in. deep, and will hold 12-13 tons of lead. At one end is a tap-
hole through which the purified lead-silver alloy is drawn off. Tb
time required for complete purification depends upon the proporticgi
of metallic impurities in the lead, and consequently varies. At the
temperature wnen lead melts ^626^ F.), the copper, iron and zinc will
readily separate and float on tne surface as a pasty mass. AntimcDj
and arsenic require a higher degree of heat, and sometimes the ^|^
cation of a blast to cause them to oxidise or volatilise. The iMat
must not be increased unnecessarily, as in proportion to it is the kas
of lead. The dross has to be skimmed off frequently to permit free
access of air to the surface of the metaL To assist in the sepan^
of mechanical impurities, wood shavings, dry leaves, or brukh wood
are mixed with the lead by stirring, when a development of gas*
takes place which causes them to rise to the surface.

Liquation is sometimes used either by itself or supplementary tn
the operation just described. It is conducted in a reverberatory for-
nace, the bed of which slopes steeply from the fire bridge to the flee
end, where it ends in a gutter leading to the lead pot. The b^*
bullion to be softened is piled on the highest part and exposed to s
temperature near the melting point of lead, when that metal «ill
gradually separate and run down the incline into the lead pot, wlu>
the metallic admixtures, having a higher smelting point, are le^
behind.

Desilverisation is effected by cupelling (rapid oxidation of the l»l
without affecting the silver), by crystallisation (Pattinson's procesl
and by the action of zinc (Parkes* process).

The English cupelling furnace is a reverberatory with morahfc
bottom, Fig. 152. The frame, of iron, and strongly bound by l^ia.
X i in. iron bands, is about 4 ft. wide in front, 3 ft. in rear, and 5^-
6 in. deep, and is filled with moistened bone ash firmly pressed is.

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METALLIFEROUS MINERALS. 543

^fter wHich a cavity is scooped out, leaving a lining 2 in. thick at the
im, increasing to 3 in. at the bottom. Lito the front end, 3 holes
kre drilled^ to serve in succession as outlets for the litharge during
lie operation. After the first one of these has become too much cor-
-oded to be any longer serviceable, it is closed, and one of the remain-
ng two is opened. A fire bridge a 14-18 in. high separates the test h
irom the fireplace c ; the
'umes and products of com-
bustion escape through two
>pening8 d into the main
lue k To prevent the test
Tom cracking, it is necessaiy
x> heat it gradually and cau-
dously to a bright red, when
part of the charge, pre-
riously melted in the iron
pot e, is introduced through
the gutter /. At first the
metal bec<)mes covered with

\ grey dross, which melts ^'ig« 156. — Emgusu Cupbl.

nrhen the temperature in-
creases ; then the blast is turned on through the nozzle ^, and forces
the litharge towards the front, where it escapes and falls into a
shallow cast-iron pot running on wheels. Fuel is added through the
ioor A, while the one at t is used for watching the operation. A flue
oarriee off the fumes collected by a hood. In cases where the lead is
introduced into the test without previous melting, openings are pro-
rided in the back wall near the tuyer through which the pigs may be
charged. In proportion as the metal in the test diminishes, fresh
lead is added, ^hen the charge has become sufficiently enriched to
tender its transfer to the refining furnace desirable, a hole is drilled
into the bottom of the test and the alloy is tapped into a pot placed
on wheels, after which the hole is plugged up and a new charge is
introduced. The final refining is conducted in a furnace of similar
construction, in which the enriched alloy is treated as above until
the LiSt traces of lead have become oxidised, and the brightening of
the silver indicates the termination of the process ; then the blast is
turned off, the fire is raked out, the silver button is allowed to set,
and the frame containing it is lowered into a small car and wheeled
away to cool off. About 4-6 cwt. bullion per hour are thus refined,
with a loss of 7 per cent., and a fuel consumption of 6-7 cwt. coal per
ton of lead.

In America, the English cupel is in universal use, but has under-
gone changes in construction and manner of operating. The test
has been in some cases increased to 6 ft. by 3 ft. 8 in. ; the original
wrought-iron hoop has been in many instances replaced by a cast-
iron ring, or, if wrought^iron has been retained, either water-cooled
coils have been added to counteract the corrosion of the litharge, or
the ring has been replaced on 3 sides by wrought-iron jackets and at
the front by cast-iron jackets of different constructions. The support
v>f the test has in many cases been made so that during the process it



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544 ECONOMIC MINING.

can be moved up and down and fddeways. Tbe filling material liu
been replaced by a limestone-day mixture, by pure cement, or by i
mixture of coarsely ground firebrick and cement. The workmg
is now generally divided into two operations: conoentratiDg &e
retort-bullion to 60-80 per cent, silver in one fdmace, and fimiihiTig
the operation, including the fining of the silver, in another. By tlm
means the concentration in water-jacketed tests, an easy operatioii
has been made continuous; the finishing, requiring special skill
being done only at intervals.

The Pattinson process is based on the fact that when argenti-
ferous lead is cooled from a molten condition with constant agitatioo,
the lead has a tendency to separate in crystals from the silver, whidi
latter thus enriches the portion remaining fluid. The plant emplopl
in the operation consists of a series (10-12) of iron pots, about 5| fi
diam. by 2^ ft. deep. At one end of the row is a pot having aboot
two-thirds the capacity of the others, called the *' market pot." Bach
pot is provided with a separate fireplace, and heated by a drcnkr
flue ending in a main flue running under the level of the floor parallel
with the line of pots. Each pot holds 6-10 tons of bullion, tb«
charging being done by cranes. Supposing the lead about to be de-
silverised to contain 20 oz. silver per ton, the bullion is placed in pot
No. 6, where it is melted ; the " pot dross," or covering of oxide whicb
forms on it, is skimmed ofl*, and the fire is withdrawn. To assist tbe
cooling, water is sprinkled on the lead, and while the temperature is
gradually decreasing, it is constantly stirred with an iron rod, thos
causing the formation of crystals. These are removed with a per-
forated ladle and allowed to drain into the pot whence they havebea
taken, after which they are placed in pot No. 6. This is continued until
i) f * o^ f o^ ^^0 contents of pot No. 6 have been transferred to Ko o,
according to whether the method adopted is by eighths, quarters, (^
thirds. The lead remaining in No. 6 will contain about 40 os. 8il?er
per ton, and is ladled into pot No. 7, while the crystals, transferred tn
No. 5, will only contain about 10 oz. A fresh charge of lead beinf
worked in No. 6, the crystals are again passed to No. 5 and tk
enriched " bottoms " to No. 7. Each pot, as it becomes filled br
crystals from the one side or by bottoms from the other, is in iH
turn crystallised. In this way the crystals, as they approach th?
market pot, become gradually poorer in silver, while the ^
bottoms, passing in the contrary direction, increase in richness. T^
various pots in the series may, from time to time, receive lead yiddine
the same amount of silver as the metal which they severally o(»tain.
During these operations a quantity of oxide is produced, and when
the charge in each pot is melted down, it is always carefallT
skimmed before cooling. The amount of dross from working W
containing 20 oz. silver per ton may be estimated at 25 per cent of
its weight. The enrichment attains its limit when 700 oz. pv [email protected]
are reached, and further concentration by these means becomes impo^
sible. The lead in the market pot should not contain more than \^
silver per ton.

Modifications of Pattinson's process are the Laveissidre, in wW^
the stirring is effected by wheels instead of by men wielding ir^



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METALLIFEROUS MINERALS. 545

ods ; and the Marseilles, which relies on steam blown in for agitating
(orposeR.

Pattinson's prooess cannot generally compete with the zinc method,
mt it has sonrived at Freiberg in consequence of certain special
lecnliarities in the lead there smelted: (a) it contains too much
opper, nickel, cobalt, tin, antimony, arsenic, and bismnth to be fit for
he zinc process without preliminary liqnation and softening ; (6) the
ilver contents being mostly large (60-1 20 oz. per ton), the ooncentra-
ion for the refinery is nearly as rapidly effected by one method afr
ly the other ; (c) production of flake litharge, for which there is con-



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