Charles George Warnford Lock.

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

. (page 62 of 76)
Online LibraryCharles George Warnford LockEconomic mining: a practical handbook for the miner, the metallurgist and ... → online text (page 62 of 76)
Font size
QR-code for this ebook


iderable commercial demand, is notably diminished when zinc is
Bed ; (J) the recovery of bismuth from the lead, which is one of the
»rofitable operations of the works, can only be effected by the older
irocess. A combination of the two methods has been developed
rhich presents notable advantages over crystallising alone. I^ad
ontaining 33 oz. of silver and upwards, after previous liquation and
of^ening when necessary, is crystallised by the method of thirds into
ioh and poor portions, tne former with 650 oz. going to the refinery,
rhile the latter, contidnine at most 33 oz., and practically freed from
ismuth, is finally desilvensed by zinc. This is generally effected by
hree additions, each of which occupies about 5 hours, and as a similar
ime is required for melting down and cleaning the surface of the
Bad, the total period of working one charge is about 20 hours. The
barge, consistiDg of 20 tons, of lead requires about 4 owt. or 1 per
ent. of its weight of zinc for complete desilverisation ; 220 lb. being
ned in the first addition, 165 lb. in the second, and 88 lb. in the third.
Lbout one-half of this quantity is subsequently recovered in the dis-
illation of the rich lead-zinc and silver alloy. The zinc removed
Iter the first addition is sufficiently rich fur the liquation pot ; but
hat of the second and third additions is put aside to be used in sub-
equent operations with fresh quantities of lead. By liquation, the
ormer product is divided into rich scum for distillation, and argent-
Gorous zinc-lead, which, together with the poorer zinc skimmings, ia
eturaed to the desilverising process. The removal of the zinc from
he desilvensed lead is effected in the softening furnace at a red heat,
oddation being achieved either by a blast of air or by the chimney
Iraught alone. The crust first formed on the surface, consisting of
inc and lead oxides, must be drawn with a rabble, but the proportion
A Uie latter gradually increases until pure litharge is produced. The
efining operation lasts 9 hours ; the proportion of zinc in the lead
liminishes from '75 per cent, at starting to *16 per cent, in 3 hours,
01 in 5 hours, -0008 in 7 hours, and -0002 in 9 hours. The latter
[nanti^ tx>rre6ponds to 1*3 dwt. per ton. When the refining is
iiushedu the softened lead is tapped into a cast-iron pot, from which,
kfter skimming the final dross, it is cast into moulds for sale. The
ich scum in which the precious metals are concentrated ccmtains in
addition to 4*05 per cent, silver, and *0153 per cent, gold, 53 per cent,
ead, 40 zinc, and somewhat more than 2\ copper. The zinc is removed
yj distillation in a plumbago crucible with a domed head and lateral
liBcharge pipe. The condenser is a cast-iron box about 20 in. high,
>f rectangular horizontal section, diminishing upwards, the capacity

2 N

Digitized by VjOOQIC



546 ECONOMIC MINING.

being abont \ oulx ft. About 4j^ owt. bf the alloy with 1 per otel
oharooal powder is charged upon a layer of lumps of charcoal place
At the bottom of the crucible. l/Hien the latter is filled, the dome i
Inted with clay, and the space inside is packed with a further qnantit
of allo^ introduced through the hole upon which the discharge pip
is finally adjusted. The crucible is placed on a square wind furnao
heated with coke, whidi requires to be twice filled with fire^ fiu
during the period of distillation, lasting 8-9 hours. When oompletet
the zinc is oollected in a lump of the shape of the oondenBer, whil
the lead remains in the crucible, covered by the unconsumed charocH
and any unreduced dross. This, when cleaned by skimming with
colander, is ladled out and passed to the refinery. The cost of woil
ing lead with 300 oz. silver per ton by the combined process is abon
13«. per ton, or about 18 per cent, lesd than by the simple Pattinsa
process. With poorer lead (150 oz. to the ton) the saving is larga
and amounts to about 21^ per cent. ; moreover the precious metals ai
rather more completely recovered.

The zinc process (invented by Parkes, but since modified in mai^
ways) is based on the observation that when molten zinc-lead is ooole
slowly, the zinc solidifies first, in a crust and carries nearly all tJi
silver with it. In operation, the silver-lead is melted in a cast-iiu
kettle, 3-3j^ ft. deep, and of a diameter to hold the contents (15-d
tons) ; some American kettles hold up to 60 tons, but in that cm
they are oblong with roimded ends. Heat is raised till zino fusee i^
the bath. The latter metal is added in 3 successive portions of |> j
and <,V< After the first addition the fire is kept up and the mass i
well stirred with a perforated ladle for \ hour, wnen the fire is reduced
and the kettle is allowed to cool. When the zinc crust is firm enoog)
it is removed with all particles adhering to the sides of the kettle
and the surface is skimmed till the lead commences to orystalliM
when the heating is renewed, and the second instalment of zinc i
added, stirring and skimming being repeated as before. Finally, th
third and last quantum of zinc goes in, and the treatment is ooncludec
The quantity of zinc is proportioned to the silver in the base bullioi
thus :—

9 oz. silyor per ton require 1) per cent zino.
18 H t, H >•

86 „ n U n

5* » w Ij „

108 „ „ 2 ,v

1" »♦ >» *t t»

A percentage of lead accompanies the silver in the zinc cmi
this is recovered by liquation in two iron pots, one of which is pla4
higher than the other and is connected with it by a pipe oast on!
bottom. The zinc skimmings are strongly heated in. the upper ii
and the liquated lead fiows into the lower one through the pipe, w£
the argentiferous zinc remains behind. The lead carries with it pi
of the silver and zinc, which, after cooling, is skimmed off. 1
liquated and purified lead is put with the original metal bef<»ei
last addition of zinc.

The output of metal by the Parkes process is good : silver, dl



Digitized by



Google



METALLIFEROUS MINERALS. r547

nnder 99^ per cent. ; gold, 98-^100 per cent. ; lead, 99-99J per cent
The cost of operating in America is about 20-25«. per ton of base
bullion. At Broken Hill the cost is 35-40«. per ton of bullion, and
the working losses amonnt to 4 per cent, lead and 2 oz. silver per ton
of bullion. Practically all the gold, 3 • 4 dwt. per ton of bullion, is
caught.

Flach's modification is conducted in 3 cast-iron pots, set in brick-
work at a conTcnient height above the floor, and heated by separate
fireplaces ; 2 hold about 6 tons each, while the third has a capacity
[>f 20 tons. Desilverising is conducted in the larger pot, and the
argentiferous zinc crust is removed to one of the smaller pots by
means of perforated ladles. When one pot has become full, it is
mbjected to liquation, and the other one serves as a receptacle for the
skimmings. The liquated lead is added to the metal in the de-
dlverising pot at the same period as in the former case. The argent-
iferous alloy is in both cases smelted in a blast furnace to separate
he last particles of lead, which is Anally cupelled. The lead remain-
Dg in the larger vessel is Icbdled into the pan of an improving furnace,
md kept at a red heat for about 12 hours, during which it is fre
luently skimmed ; at the expiration of this period it is cast in moulds,
hnd forms market lead.

In Corduri^'s method the lead is brought to a red heat, and super-
heated steam is forced through it, when the oxygen contained in the
atter causes the zinc to oxidise, while the lead is but slijerhtly affected ;
he zinc oxide rises to the surface and is skimmed off. The zinc^silver
Hoy may be treated in the same way, when the zinc will oxidise and
eparate from the argentiferous lecbd alloy. The latter is finally
upolled to obtain pure silver and litharge.

By another way, the alloy is kept at a moderate temperature under
cover of chloride of lead for about 24 hours and continually stirred,
'hen the metallic zinc is converted into zinc chloride and the lead
bloride into metallic lead.

Schnabel's process consists in the digestion of the argentiferous
inc and lead oxides with a hot solution of carbonate of ammonia
nder pressure in gas-tight vessels. The zinc oxide dissolves and is
mverted into zinc carbonate, and a silver-lead alloy is obtained in
buitable state for refining. The ammoniacal solution is distilled to
hoover ammonia, and the basic zinc carbonate is converted by calci-
ation into the oxide, which is used as paint.

Balbach and Faber du Faur employ a retort furnace as shown in
ig. 153 : a, fireplace ; b, grate ; c, ^el door ; d, retort ; e, flue. It is
red with charcoal or coke, and when the retort is red-hot the charge
\ introduced, consisting of a mixture of finely broken zinc crust and
larcoal smalls, and varying according to the size of tlie furnace from
50 to 400 lb. zinc crust, and 3 to 5 lb. charcoal. Then a condenser is
laced over the mouth of the retort, and the temperature is at once
kised to white heat. Should it be neglected to maintain this hiich
tmperature uniformly, a crust of chilled alloy will form on the metal,
nder which zinc fumes accumulate, causing an explosion if the
»mperature is once more raised. The metallic zinc collects in the
mdenser, from whence it is from time to time removed, remelted,

2 N 2

Digitized by LjOOQIC



548



ECONOMIC MINING,





5


s


i

5


"^

E


€b


-






H


*<%*





» «


♦ V*






1




1

-1-




I


1


1 1


1


1-


1


1 1


1


1




1 1


1




_


,



Fio. 157.— RDnriHO Rktobt.



and oast into thin platee, to be used again in desilTerising. In tiiif
way 60-80 per oent. of the fine is recovered, 40-50 per cent, in Ibi
metallic state, and 20-30 per cent, as oxide. The latter collects aroonl
the month of the condenser ; it is scraped off, packed into soitaldi

vesselB, and taken to zinc works €■
reduction. The argentiferooB lead ?<»
mainin^ in the retort is tapped ofi^ cari
into thin plates, and capelled in «i
English furnace. The entire opera^
requires 8-10 hours, according to tibi
percentage of zinc in the alloy. Tki
romace has found most favour amoo|
zinc distillers, because it is easy to keaf
at a uniformly high heat, and tb
retort can be quickly emptied, deand
and refilled for a fresh operation, vai
endures longer. In America the wai
has been increased up to 1000 lb.

In the R68sler-£delmann
operated at Hoboken, near Antwerp, i
addition of a little aluminium to the j
produces a rich zinc-silver crust, free from lead, and avoids prodi
of bulky scums. An alloy of * 5 per cent, aluminium with the zinc do
not oxidise at desilverising temperature ; thus desilverisation is m
possible at one operation, without much stirring, and with much ]
zinc. In practice, the zinc-aluminium alloy, previously prepared, h
thrown upon the 1^ bath, when the latter has acquired the necesBafj
temperature, varying with the contents in silver, but about 750^^
900^ F. Then the whole is stirred and allowed to cool, wbereapoB
the molten lead, which at the low temperature is no longer capabh
of holding the zinc, gives it up again. The free zinc, having in tb<
meanwhile taken up Qie silver, rises to the surface of the bath, wheiM
it, together with some lead, is ladled off. In order to get rid of the
excess of lead, the alloy is charged into a cast-iron pot with an outk^
at the bottom, and slowly heated, liquating and drawing off the gr^ti
part Subsequently the temperature is raised to red-heat, for me^ '
the zinc-silver alloy, as well as for separating it from the rema
of lead present, the former floating on top of the latter, whence it i^
ladled, care being taken not to touch the lead underneath, whidi \
drawn off afterwards. The zincndlver alloy consists of 20-40 pi
cent silver, according to the richness of the silver-lead treated, 5 pei
cent, lead, 2-4 per cent, copper, and 60-70 per cent zinc. It amounti
to about 2 per oent. of the silver-lead treated, while by the old pn>
cess about 15 per oent. zinc scum, consisting of 4-6 per cent silver
70^0 per cent lead, *5 per cent copner and 10 per cent zinc wai

Produced. For working up the zinc-suver alloy there are two wajt
^he first is to treat the granulated alloy by hydrochloric or dilate
sxdphuric acid, whereby the zinc is got as a salt, and the silverin Um
«hape of slime. The second way is by electrolysis, whereby the spdtei
obtained as a metal of high puri^, consisting of * 0099- * 0044 ptf
t iron, •0114-'0210 per oent copper, •0341-*0600 per oent lc»d,



Digitized by



Google



METALLIFEROUS MINERALS. 549

from a trace to '0020 per oent diver, and 99*9446-99*9226 per oent.
cine. This metal, of coarse, commands a mnch higher price inan that
>f ordinary spelter, the gain nearly covering the cost of electrolysis,
rhe electrolyte consists of a solution of zinc chloride in magnesium
ddoride. Its sp. gr. is about 1 -2 to 1 -27. The cathodes are vertical
nrcular sheets of metallic zinc fixed upon a horizontal spindle, the
tatter revolving just above the surface of the bath. The spelter is
thereby obtain^ in sheets. The residue of the anodes, got in the
ihape of slime, after the electrolytic extraction of the zinc, contains
ibout 75 per cent, silver and 12 per cent. lead. A small quantity of
3hloride of silver is also formed. The oxides of copper, zinc and iron
ire dissolved in very dilute n2S04, while the chloride of silver is
reduced at the same time to the metallic state by iron shavings. The
diver slime now contains nearly 15 per cent, lead, some copper, and
30-85 per cent, silver. It is smelted upon a cupel, whereby the
remainder of the lead is oxidised and separated as litharee. About
150 lb. silver slime are refined in 8 hours, and it is possime to refine
I charges in 24 hours. Cupellation is done away with, and with it
Bie reviving of litharge and other bye-products. In lieu of it there
is only the short refining process on the cupel. As there is only a
very small quantity of litharge produced, practically the whole of
the silver-lead is worked at once into refined lead, so that no
Ribsequent desilverisation of the revived bullion, as hitherto, is
required.

Utilising Heat of Slag. — Careful measurements * at Broken Hill,
irith a view of determining the true values of the heat wasted in the
dags, show that, although a comparatively small proportion only of
the total heat pit)duced by the combustion of the coke supplied to the
famaces is contained in the slags (the remainder being found in the
chemical reactions of the ores and fluxes in the furnace or passing up
the flues with the gases), the amoimt is still large enough to be of
eonsiderable importance, and its utilisation for steaming purposes
capable of effecting an important economy. The following are the
principal data determined by the measurements : sp. gr., 3*8; tem-
perature of exit from furnace (average), 2000° F. ; specific heat, 25 ;
latent heat per lb. (probably), 120 heat-units; total heat per lb.,
620 heat-units ; average output of slag from one furnace, 112 in. by
60 in., 4400 lb. per hour ; theoretical mechanical equivalent of slag
from one furnace, 1064 h.p. As in the case of steam production from
^e combustion of coal, only a small portion of this waste power can,
by any known method, be utilised for mechanical work. Direct con-
tact methods were discarded on trial for a method of imparting the
heat to the water through the medium of a metallic casing* the prac-
tical result being that with a comparatively small and loexpensive
boiler, the whole output of slag from one 112 x 60 in. silver-lead
fomace can be readily utilised, with a production of over 60 h.p.

Fumaee Bye-Produets. — (a) The term ** speiss " is applied to sub-
stance consisting essentially of iron in combination with antimony or
arsenic, containing besides varying proportions of valuable metals, as
well as some sulphur, lime, silica, copper, lead, molybdenum, zinc, &o.
* J. Howell and E. A. Ashcroft, Trans. AoBtr. Inat. Min. Engs.

Digitized by VjOOQIC



S50 ECONOMIC MINING.

A praotical attempt to utilise speifis at the Richmond works, Keyada,
consisted in tapping the molten speiss into pots having a lining of
clay and limestone, and at the same time adding a sufficiency of lesd
or litharge to collect the precious metals. The carbonic acid set firee
hy the heat of the molten speiss serves to keep the contents of the pot
in ebullition, so that the lead gradually sinks to the bottom of the
vessel, carrying the precious metals with it. As a result^ there was
extracted about 67 ^r cent, of the value. This process was improved
upon by L. W. Davies, who adds about 25 per cent, molten lead to the
molten speiss, in a metal converter under an air pressure of 17 lb.
The converter is cylindrical, and has a lining of 2^ in. of firelnrick.
The economic results obtained have been favourable : the percentage
of silver extracted is reported at 83*50, and of gold 89*28.

(6) " Matte " is chiefly compounded of iron and sulphur, eoine
lead, copper, gold, and silver being also usually entrapped in it. It is
re-roasted (to remove sulphur) and re-smelted, the copper beo6miii^
concentrated thereby and carrying the precious metals.

(c) ** Slag " contains the gangue and waste matters of the ores an^
fluxes, and should not afford more than 1 per cent, lead or 1 ok. ralvei
lYOT ton. Bich slags are i-e-smelted. At most works, each potful d
slag is separately hand-picked. As soon as the pot is cool, its content!
are deposited on the 8U]:%su>e of the dump, where, when quite cold, tlM
cone is bioken up and carefully examined for a cake of matte, whicli
is generally found at the point of almost every cone, in addition U
which there is often a smaller cake of speiss below the matte, and
sometimes a button of lead at the extreme point of the cone. Th«
matte contains practically all the copper contents of the original ore,
besides a considerable amount of silver ; but the speiss is generall^^
very poor in precious metals, and is generally thrown away.

A very small quantity of zinc renders both the matte and slag sc
pasty that perfect separation is impossible ; in that case, each po<
of slag is allowed to cool for 10 minutes after being filled, or until s
solid '* shell " has been formed j-1 in. thick all round against the iron ;
the still liquid interior is then ^ured, either by pricking the to^
crust or by piercing a clay plug m the bottom. Practically aU thi
suspended globules of matte collect in the *' shell," which weighs
10 per cent, of the whole potful ; this is re-smelted.

At Leadville, Colorado, a separate furnace is employed for re^
smelting * slag, being, in the case described, merely an old 86 by 80 in.
lead-furnace, having itscruqible filled with well-tamped sand covered
by a course of firebricka The rich slag averaged 5*3 os. silver pei
ton, and 8107 tons were smelted in one month, together with 402 ton^
pyritous copper ores averaging 10 per cent, copper and 11 oz- sihreil
per toiL The matte produced averaged 20 per cent, copper asil
93 * 3 oz. silver per ton, while the clean slag averaged only f oz. pezj
ton ; the loss of copper was insignificant, averaging only ^ per oentj
or less. The fuel-consumption was 1 to 10 of charge (as against 1 to 5|
for the original ore-smelting), and the total expense per ton of material
smelted was only 6«.

Flue dust and Fume. — ^Flue dust consists essentially of fine partid^s:
* H. A. KeUer, Trans. Amer. Inat Min. Engs., 1892.

Digitized by VjOOQIC



METALLIFEROUS MINERALS. 551

of ore and fael carried over by the blast, while fame is lead oxide and
sulphate deposited on cooling. The two are of necessity intimately
mixed and must be dealt with together. The material often aggror
gates 2J-3 per cent, of the dry ore charged. Its composition is never
definite. While containing much partially oxidised lead, there are sure
to be also many impurities, notably zinc, arsenic, and antimony. The
essential element of condensation or collection of the fame is a very
long flue, in the construction of which it is well to make provision for
retardation of the current of gases through it on their way to the
chimney. Such provision may be made by giving a large capacity to
the flue; by giving it an angular or zigzag course; by interpolating
large chambers, which are best placed at the far end of the flue, or at
dgzags in its course ; by introducing baulks or mid-feathers here and
there, against which the current may impinge ; or by hanging within
the flue such things as iron hooping, bushes, old nets, &c., upon which
he solid matters may be deposited. These various contrivances are
idopted at difl'erent works. But while the solid element of the fame
an be arrested in this way, long flues have no influence in arresting
he escape of sulphurous add from the chimney, and ordinary modes
f washing with water &il to remove more than a moderate proportion
f it. Tlus sulphurous acid creates a serious muisance, and its escape
tnutilised is a considerable loss. By Wilson and French's method,
he oooled famace gases are forced through water in such a way that
he water and fume are brought into very close contact and thoroughly
lixed ; the solid element is thus effectually wetted and retained in
he water, and the soluble gases are, as far as the dissolving power of
be water permits, dissolved. The solid matters have to go back to
be smelting famace, but are always troublesome to treat. Moistening
( only a temporary expedient for making it cake into a solid fomu
. better procedure is to mould it into bricks with milk of lime ; or,
Btter stiU, if obtainable at a moderate price, to bind it together with
solution of iron sulphate. When thus compacted, it may be added
» the charge and treated as ore.*

New Processes for Zinc-Lead Sulphides. — The flue dust and fame
ifficullnr is enormously increased in the case of zinc-lead sulphide
66, ^which are daily becoming of more importance as the mines
aoh greater depths and the (more or less) oxidised surface ores are
Jiansted. Enormous quantities of such ores are now exposed, and
ntain notable proportions of the precious metals, but cannot find a
ady market, because their zinc contents occasion great loss and iur
Dvenienoe in ordinary smelting operations : if it be attempted to
IX off the zinc, volatile metal rises to the throat of the stack and is
ere oxidised, forming hard, infusible lumps, and compelling frequent
[>ppage8, whQe the great heat necessitated in the lower part of the
maoe involves considerable loss of lead and silver by volatilisation ;
the mixed ore be smelted for zinc, the associated lead forms a fusible
m pound with the silicious materials of the retorts, and precludes
coetsaful results. Something may be done by mechanical separation
the two ores, blende and galena, as already described (p. 521) ; but

* See T. Egleitton, '* Collecting Flue-dost at Enn," Ttans. Amer. In&i. Min.
1,^1883.

Digitized by VjOOQIC



552 ECONOMIC MINING.

when both are argentiferous and fine grained, such a method is im-
practicable. Modified smelting, and chemical treatment (wet) in
several forms have attained some success.

(a) In the Lewis-Bartlett process,* the smelting aims directly ak
causing volatilisation of the zinc, special means being adopted for
catching the zinc oxide and lead sulphate fumes, while the residiul
matte and slag contain the iron, some lead, and most of the silver,
and are fit for charging the ordinary smelting furnace. The oc^ected
fumes are sold as white pigment.

At the works of the richer Lead Co., Joplin, Missouri, the prooes
has long been used on rich galena. The ore is first roasted in s modi-
fied Scotch hearth (called a Jumbo or Mofifet), producing in pig-kad
about 60 per cent, of the lead in the ore, and a large quantity of kad
slag and lead fumes. The Jumbo or Moffet ore heartn has a lioUoir
cast-iron back perforated with 5 or 6 holes, the single baok serving
for both sides of the furnace. The lower part is a dam of oast iron,
which reaches to the bottom of the basin, except where it is cut away
at one comer, to allow the metal to be drawn off from one side. Being
hollow, this dam is always filled with molten metal. The air supplied
to this hollow back is made to pass through the side walls of thi
hearth, in order to heat it. The basin is only 8 in. deep, and the lead
running out of the overflow is drawn from the bottom of the batii by
a tapping pipe reaching nearly to the bottom of the basin. The ore
is fed on to the hearth, mixed with coal and lime : to 1 ton of ore,



Online LibraryCharles George Warnford LockEconomic mining: a practical handbook for the miner, the metallurgist and ... → online text (page 62 of 76)