GIFT OF
HYDROMETALLURGY
OF SILVER
WITH SPECIAL REFERENCE TO
CHLORIDIZING ROASTING OF SILVER ORES AND THE
EXTRACTION OF SILVER BY HYPOSULPHITE
AND CYANIDE SOLUTIONS
BY
OTTOKAR HOFMANN
Mining and Metallurgical Engineer, Member of the American Institute
of Mining Engineers, of the American Chemical Society
and of the American Electrochemical Society
NEW YORK AND LONDON
HILL PUBLISHING COMPANY
1907
T~/V
Copyright, 1907, BY HILL PUBLISHING COMPANY.
ALSO ENTERED AT STATIONERS* HALL, LONDON, ENGLAND
All rights reserved
. , . », «,„,,,
: -:.:• '•
Hill Publishing Company New York U.S.A.
PREFACE
THE silver ores which are treated by a hydrometallurgical
process are mostly complex sulphide ores consisting of quite a
number of different metal sulphides. In order to render soluble
the silver contained therein the ore is roasted with an addition
of salt (chloridizing roasting), by which process the silver is
converted into silver chloride. In this chemical reaction, how-
ever, all or nearly all the constituent minerals of the ore partici-
pate, which makes the process rather complicated and we may
even say delicate, because the formed metal chlorides are volatile
and induce the silver chloride to volatilize too, and in order to
keep this loss at the minimum great care and skill are required.
The solvent, whether this be sodium hyposulphite or concen-
trated brine, will extract all the silver which was converted into
chloride, and the final result of extraction depends, therefore,
entirely on the quality of the roasting. This being the case, it is
obvious that chloridizing roasting is the most important part of
the process, and that a metallurgist can not expect to be success-
ful in the hydrometallurgy of silver unless he has a thorough
knowledge of chloridizing roasting and the ability to apply
skilfully this knowledge in actual practice. For this reason a
large part of this treatise is devoted to the art of chloridizing
roasting, which I consider to be especially necessary, as there is
no new literature on this subject, though great advance has been
made in it.
Chloridizing roasting was originally studied and practised in
relation to amalgamation. In amalgamation not only the silver
amalgamates, but the base-metal chlorides also amalgamate,
which greatly interferes with the process, causing a poor extrac-
tion, a great loss of silver and mercury, and the production of
very base bullion. To avoid these difficulties and to make the
process applicable to a greater variety of ores, these objectionable
metal chlorides are partly expelled and partly converted into
iii
337632
iv PREFACE
oxides by increasing the temperature of roasting. The expulsion
of the volatile metal chlorides, however, induces quite a percent-
age of the otherwise not volatile silver chloride to volatilise,
thus causing a considerable loss of silver.
I was the first to introduce the process of lixiviation with
sodium hyposulphite in Mexico, in 1868, and made a special
study in actual practice of chloridizing roasting, and in course
of time, and as experience was gained with different ores, became
convinced that chloridizing roasting as practised for amalga-
mation was not the proper way to roast for lixiviation. A large
percentage of base-metal chlorides in the roasted ore is detrimental
to a successful extraction of the silver by amalgamation, while
in lixiviation they do not interfere with the extraction; why then
should we expel and destroy these chlorides by increased heat
at a sacrifice of silver? The expulsion of the volatile compounds
by increased heat is the sole cause of the silver loss by volatiliza-
tion, and if we avoid this we consequently will avoid this losSj or
at least reduce it to the minimum. I therefore modified the mode
of chloridizing roasting, inasmuch as, instead of expelling the
volatile compounds by increased heat, I endeavored to retain
them as much as possible in the roasted ore by using the lowest
permissible temperature — a temperature sufficiently high to pro-
duce the chemical reaction but not high enough to expel the metal
chlorides. Of course such a condition cannot be maintained with
theoretical exactness, but I succeeded in greatly reducing the
loss of silver by volatilization even with ores rich in arsenic.
By this modification in chloridizing roasting a marked step forward
was made in the hydrometallurgy of silver.
The second part of this treatise is devoted to the extraction
of the silver from the roasted ore by different solvents. The last
chapter deals with the cyaniding of silver ores. This process, so
extensively and successfully used for the extraction of gold, is
still more or less in its experimental stage with regard to silver
ores. However, such very promising results have been obtained
with certain ores, that further experiments and a thorough
investigation of this subject are to be recommended. For complex
auriferous silver ores a combination of the sodium hyposulphite
and the cyanide processes is most promising.
By far the larger part of this treatise is a record of my long
years of experience, studies, and experiments on a commercial
PREFACE V
scale rather than the product of compilation, and it will be found
that attention is paid to many things apparently unimportant
but which, in actual practice, I have found to be of great impor-
tance, and on the correct execution of which the success of the
enterprise may often depend.
I hope that this treatise will prove a friend and adviser to the
student of hydrometallurgy as well as to the metallurgist in active
service in the field.
OTTOKAR HOFMANN.
KANSAS CITY, Mo., January, 1907.
TABLE OF CONTENTS
PREFACE .
TABLE OF CONTENTS
LIST OF ILLUSTRATIONS
PAGES
iii
vii
ix
PART I. CHLORIDIZING ROASTING OF SILVER ORES
CHAPTER I. THEORY OF CHLORIDIZING ROASTING 1-10
Behavior of different minerals in chloridizing roasting, 5.
Classification of ores in relation to chloridizing roasting, 9.
CHAPTER II. CRUSHING OF THE ORE 'r 11-14
CHAPTER III. PERCENTAGE OF SALT REQUIRED . . .; . . 15-19
The proper time to add the salt, 16.
CHAPTER IV. Loss OF SILVER BY VOLATILIZATION .... 20-25
Method of ascertaining the loss of silver by volatilization, 22.
CHAPTER V. METHODS OF ROASTING 26-41
Chloridizing self-roasting, 26. Chloridizing heap-roasting,
27. Chloridizing roasting with steam, 31. Chloridizing roast-
ing of silver ores containing gold, 35. Chloridizing roasting
for amalgamation, 37.
CHAPTER VI. CONSUMPTION OF FUEL . . 42-44
CHAPTER VII. REVERBERATORY FURNACES WORKED BY HAND . 45-61
The single-hearth reverberatory, 45. The two-story single-
hearth furnace, 46. The long reverberatory furnace, 47. The
two-story long furnace, 60.
CHAPTER VIII. MECHANICAL ROASTING FURNACES .... 62-86
Mechanical furnaces fed by charges, 63. Mechanical roasting
furnaces with continuous feeding, 71.
CHAPTER IX. COLLECTING THE FLUE-DUST 87-93
CHAPTER X. SULPHATING ROASTING 94-98
CHAPTER XI. CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 99-126
Roasting experiments, 101. Roasting in the Stetefeldt fur-
nace, 104. Re roasting the ore from the shaft, 106. Re roasting
the ore of the Stetefeldt furnace in the modified Howell furnace,
109. Application of steam, 110. Conclusions, 111. Roasting
in the White-Howell furnace, 112. Roasting in the modified
Howell furnace, 113. Additional chlorination after the ore
has left the furnace, 115. Results, 115. Loss of silver by
volatilization, 117. The roasted ore, 118. Consumption of
wood, 119. Cost of roasting in the modified Howell furnace,
Vlll
TABLE OF CONTENTS
PAGES
119. Summary, 120. Roasting in the reverberatory furnace,
121. Oxidizing roasting, 122. Treating the oxidized ore with
cupric chloride, 125. Consumption of wood in the reverbera-
tory furnace, 125. Cost of roasting in the reverberatory
furnace, 126.
CHAPTER XII. CHLORIDIZING OF CALCAREOUS ORES ....
Roasting in the Bruckner furnaces, 128. Adding the salt in
the furnace, 135. Adding the salt in the battery; self roasting,
140. Balling of the ore, 142. Roasting in the reverberatory
furnaces, 144. Conclusion, 151.
127-152
PART II. EXTRACTION OF THE SILVER
CHAPTER XIII. LIXIVIATION WITH SODIUM HYPOSULPHITE . 155-184
Base-metal leaching, 157. Silver leaching, 174.
CHAPTER XIV. PRECIPITATION OF SILVER . . ... . . 185-193
CHAPTER XV. TREATMENT OF THE PRECIPITATE 194-214
CHAPTER XVI. CONSTRUCTION OF TROUGHS . . . '. . . 215-218
CHAPTER XVII. TROUGH LIXIVIATION 219-250
The troughs, 224. Sluice-tanks and sluicing, 225. Ar-
rangement and operations, 229. Precipitating vats, 233.
Practice of trough lixiviation at Cusihuiriachic, 240. Trough
lixiviation experiments on a large scale, 243. Time required
for base-metal leaching, 244. Quantity of water required, 245.
Quantity of silver dissolved by the base-metal solution, 246.
Silver leaching, 246. Quantity of solution required, 248.
Fineness of the precipitate, 248. Advantages of trough
lixiviation, 249.
CHAPTER XVIII. THE RUSSELL AND Kiss PROCESSES . . . 251-255
The Russell process, 251. The Kiss process, 254.
CHAPTER XIX. THE AUGUSTIN PROCESS ... ... . 256-257
CHAPTER XX. EXTRACTION WITH SULPHURIC ACID . - . .. . 258-280
Extraction of silver from copper matte, 258. Extraction of
silver from black copper, 278.
CHAPTER XXI. THE ZIERVOGEL PROCESS ....].. . 281-282
CHAPTER XXII. TREATMENT OF SILVER ORES RICH IN GOLD . 283-286
CHAPTER XXIII. CYANIDATION OF AURIFEROUS SILVER ORES . 287-328
Treatment of raw ore, 287. Cyaniding auriferous silver ores
at Palmarejo, Mexico, 288. Cyaniding auriferous silver ores at
San Salvador, C. A., 321. Testing the cyanide solution for
gold and silver, 326.
INDEX 329-345
LIST OF ILLUSTRATIONS
FIGURES PAGES
1-2. Single-hearth reverberatory furnace 46
3. Two-story, single-hearth reverberatory furnace . . 47
4-6. Long reverberatory furnace . . . . . . . 49-50
7-8. Plan and elevation of working door 57
9 A. The Kiistel working door 58
9 B-9 C. Device for working door ........ 59
10-12. Long reverberatory furnace, two-story .... 60-61
13-14. Bruckner roaster . 63
15-17. Hofmann improved Bruckner furnace .... 68
18. O'Harra furnace 73
19-20. Horizontal and cross-section of Ropp furnace . . 75
21 A-21 B. Ropp furnace, longitudinal elevation and plan . . 76
22. Howell-White furnace . ...... . . . 78
23. Howell furnace, discharge end and ore-vault . . 79
24. Stetefeldt furnace 84
25. Feeding machine, Stetefeldt furnace 85
26. Vertical section of Hofmann dust collector ... 88
27-28. Details of bars and bearings, Hofmann dust collector 89
29. Position of bars, Hofmann dust collector. ... 90
30. Horizontal section, Hofmann dust collector ... 92
31. Leaching tank, vertical section . . "~7 . . . 158
32. Leaching tank, plan .......... 159
33-34. Brass clamps for \\- and 2-in. hose 161
35. Calcium sulphide plant . . . ... . . . 187
36. Distributing trough for milk of lime 188
37-39. Boiler and pressure tank for calcium sulphide . . 189-190
40-41. Air blow-off drum 200-201
42. Horizontal pressure tank, for solution .... 201
43. Cast-iron flange union for discharge pipe of pressure
tank . . . . ..- 202
44. Apparatus for the manufacture of lye . . . . . 204
45. Pressure tanks for treatment of precipitate . . . 206
46-47. Drying and roasting furnace for silver precipitate . 208-209
48. Dust-collecting afrangement for cupeling furnace ." 212
49-51. Trough: cross section, connection, union .... 217
52-54. Settling-tank arranged for sluicing 226-227
55. Wheel for closing discharge gate ...... 228
56. System for continuous trough lixiviation .... 230
57-58. Precipitation tank, vertical section and plan . . 234-235
ix
x LIST OF ILLUSTRATIONS
FIGURES PAGES
59. Precipitating vat 236
60-61. Filter frame . . 237
62-63. Lump-grinding machine, elevation and plan . . . 238-239
64-65. Lump-grinding machine, mantle and muller . . . 240
66. Stir tank, vertical section . . 264
67-69. Cast-iron pressure tank . . j . '. ,, . . . 266
70. Tower for refining cupric sulphate solutions . . . 268
71-72. Pan evaporator, longitudinal vertical section and plan 271
73. Device for discharging blue vitriol 276
74-75. Cyanide leaching plant, plan and elevation . . . 293-294
76-77. Plan and section of slime plant 306-307
78. Agitation vat . rf . . ... . . . . . . 308
79. Decantation vat ./..<... .... 311
80. Timber foundations supporting decantation vats of
slime plant 312
81. Decantation vats in course of construction . . . 312
82. General arrangement of slime plant 314
83. Three of the agitation vats and top of two of the
decantation vats . 314
PART I
CHLORIDIZING ROASTING OF SILVER ORES
THEORY OF CHLORIDIZING ROASTING
THE object of chloridizing roasting is to convert the silver in
the ore into silver chloride, in which state, while not soluble in
water, it becomes soluble in sodium hyposulphite and other solu-
tions like hot concentrated brine, potassium cyanide, etc., by
means of which it can be extracted from the ores. It is one of
the most complicated, and in the hydrometallurgy of silver the
most important, of metallurgical operations. The results of the
subsequent extraction of the silver by the solvent depend entirely
on the quality of the roasting. Silver chloride dissolves easily,
and even a very dilute solution of sodium hyposulphite will
extract all the silver which was converted into chloride during
roasting, so that it is of the greatest importance that this part of
the process be executed intelligently, and with great care and skill.
The ores which are subjected to chloridizing roasting are
usually complex sulphide ores, though in some instances ores
almost free of sulphides are roasted successfully, but these are
exceptional cases. To effect chloridizing roasting chlorine has
to be generated in the ore while being subjected to heat. This is
done by an addition of salt (sodium chloride) to the ore. But
not only the silver is converted into a chloride; all the constit-
uent parts of the ore also> undergo a change, quite frequently even
the gangue. During the first part of the roasting the sodium
chloride remains indifferent, while the metal sulphides oxidize,
forming metal sulphates and sulphurous acid; then by the action
of these sulphates on the salt rather complicated reactions take
place, by which metal chlorides, chlorine, hydrochloric acid, and
sulphurous chloride are formed.
The decomposition of the sodium chloride and the chlorination
of the silver and other metals is effected in the furnace in different
ways:
(1) In oxidizing the metal sulphides there is always, besides
3
4 HYDROMETALLURGY OF SILVER
sulphurous acid gas, some sulphuric acid gas formed. Not so
much in the beginning as later, when part of the metal sulphides,
especially the iron, have changed into oxides, which then act as
a contact substance on the sulphurous acid, converting it
into anhydrous sulphuric acid, which then decomposes the sodium
chloride. The formation of sulphuric acid increases much if a
liberal amount of air is permitted to enter the furnace, and as
the sulphuric acid plays an important part in chloridizing roasting,
provision should be made, in the construction of the furnaces,
that they may receive as much air as required.
(2) By the reaction between metal sulphates and the sodium
chloride, by which metal chlorides and sodium sulphate are formed.
This is the principal reaction for the formation of chlorides. The
metal sulphates which act most energetically in this respect are
those of iron and copper, for which reason ores containing an
ample amount of iron pyrites and some copper sulphides will be
found to chloridize the best.
(3) Besides chlorine, there is also hydrochloric acid formed,
owing to the moisture in the air and fuel. Hydrochloric acid
acts very energetically, and sometimes it is of advantage to
produce larger quantities of it, in which case steam is admitted
into the furnace to supply an extra amount of moisture.
(4) The fumes of volatilized salt (sodium chloride) act also in
chloridizing the ore. Quartz decomposes the salt, forming
s'licate of soda and chlorine, but it takes a rather high heat for
this reaction, and only in exceptional cases does it come into play.
(5) Volatile metal chlorides act also, chloridizing the silver,
whereby they are reduced to subchlorides or changed into oxides.
Cupric chloride acts very energetically in this respect.
If salt and ore are charged together, the salt is not decomposed
until the formation of sulphates begins, and the first stage in
roasting is, therefore, a mere oxidizing process. Whatever sul-
phuric acid is formed during this period by the oxidation of the
sulphides acts more readily on the base metals, forming sulphates,
than on the salt. Likewise it acts more readily on the lime and
other earthy matters of the gangue. The principal part of the
chlorination takes place by the reaction between the metal
sulphates and the salt. The oxidizing and chloridizing periods
are quite distinct, and can be easily observed by the appearance
of the ore in the furnace and by the smell of the fumes of a sample
THEORY OF CHLORIDIZING ROASTING 5
taken from the charge. During the oxidizing period the glow of
the surface of the ore is much brighter than the inside, and the
particles brought to the surface by stirring brighten instantly to
a lighter red. The fumes of a sample have a strong, choking
smell of sulphurous acid. During the chloridizing period, if an
excessive fire is not kept up, the surface of the ore assumes a very
dull red, while the deeper layers are of a brighter glow, which,
however, becomes dull shortly after the particles are brought
to the surface. The fumes of a sample have a mild but distinct
odor of chlorine. The charge swells and becomes loose and
woolly.
It was mentioned above that the ores, which are subjected to
chloridizing roasting, are mostly complex argentiferous ores, and
as the roasting is much influenced by the behavior of the con-
stituent parts of the ore and has to be modified according to
the requirements of one or the other of the constituents, a
knowledge of the behavior of the different minerals and the
changes they undergo during chloridizing roasting is therefore indis-
pensable in order to conduct the process intelligently.
BEHAVIOR OF DIFFERENT MINERALS IN
CHLORIDIZING ROASTING
Iron Pyrites. — During the oxidizing period, sulphurous and
sulphuric acids are formed, of which the former escapes entirely,
while part of the latter combines with lime and other earthy
matters of the gangue, and part combines with the iron, forming
sulphates. The iron changes into ferrous and ferric sulphates and
into ferric oxide. The ferrous and ferric sulphates act on the
salt, forming ferrous and ferric chlorides and sodium sulphate,
while some of the chlorine combines with sulphur to form sulphurous
chloride, which escapes as gas. In the course of the process
both these iron chlorides give off their chlorine, chloridizing the
silver and changing into ferric oxide. In practice the reaction
is not quite so complete, and in the finished charge we find,
besides the ferric oxide, some ferric sulphate and some ferrous
and ferric chloride.
The iron chlorides decompose easily and act as the principal
chloridizers, for which reason it is very desirable, in fact often
necessary, to have iron sulphides in the ore.
6 HYDROMETALLURGY OF SILVER
Copper pyrites consists of the sulphides of copper and iron.
During oxidizing, cupric sulphate, cuprous and cupric oxides
are formed. The cuprous oxide, however, soon changes into
cupric. During chloridizing, sulphurous chloride (which vola-
tilizes), cupric and cuprous chlorides are formed. Both these
copper salts melt below red heat, and are absorbed by the ore,
thus becoming finely divided through the ore and coming in inti-
mate contact with the silver. Both are volatile. At a higher
heat the cupric chloride gives off part of its chlorine, chloridiz-
ing the silver and changing into cuprous chloride. In presence
of steam, hydrochloric acid, cuprous and cupric oxides are
formed.
The iron sulphide of the copper pyrites undergoes the same
chemical changes as the iron pyrites. For this reason, and for
the fact that cupric chloride gives off part of its chlorine, copper
pyrites is a very good producer of chlorine during roasting.
In the roasted charge we find cupric oxide, cupric sulphate,
ferric oxide, ferric sulphate, ferrous and ferric chlorides, and
cuprous and cupric chlorides. If such a charge is subjected to a
prolonged roasting at a high heat (dead roast), all the iron as
well as the copper will be changed into oxide.
Other copper ores, like gray copper ore, fahlerz, etc., undergo
the same changes.
Galena (lead sulphide) undergoes the changes much slower.
It cakes easily, for which reason the temperature in the begin-
ning has to be kept low until most of its sulphur has been oxidized.
During oxidizing, sulphurous acid, lead oxide, and lead sulphate
are formed. The lead sulphate does not decompose the salt
at a roasting heat and, therefore, does not take an active part in
the generation of chlorine. When air has free access, most of
the lead is converted into sulphate and but little into chloride,
while, if the supply of air is limited, much more lead chloride
is formed, and thus becomes a consumer of chlorine. It is volatile,
and volatilizes without giving off any chlorine. Lead oxide is
volatile too, while the sulphate remains more indifferent. In
the roasted ore we find lead sulphate and lead chloride, but much
less of the latter.
By the above it can be seen that lead sulphide is not a desir-
able constituent part of a roasting charge.
Zinc Blende. — During the oxidizing periods zinc sulphide
THEORY OF CHLORIDIZING ROASTING 7
changes into zinc oxide and zinc sulphate, but the sulphate does
not act decomposingly on the salt. By the action of the chlorine
and hydrochloric acid zinc chloride is formed, which is very
volatile and goes off in heavy fumes, which increase when the
temperature is raised. These escaping fumes induce the silver
to volatilize, for which reason ores rich in zinc blende have to be
roasted at a low heat to avoid an excessive loss of silver.
In the roasted ore we find principally zinc oxide, then zinc
sulphate and chloride.
Zinc blende, as a rule, contains more or less iron sulphide,
some of its varieties as much as 22 and even 28 per cent. The
iron sulphide takes, of course, an active part in the generation
of chlorine; still it takes much skill to chloridize satisfactorily
the silver contained in zinc blende. This subject will be treated
exhaustively in another chapter.
Arsenical Pyrites. — This consists of arsenic sulphide and
iron sulphide. Arsenic is very volatile and begins to come off from
the ore in dense fumes right at the beginning and before other
sulphides are ignited. During this part of the process much
arsenate of silver is formed, up to 50 and 54 per cent, of the total
silver contained in the ore. This silver compound is soluble in a
solution of sodium hyposulphite. During the chloridizing period,
however, most of it is decomposed without volatilizing, if the
temperature is kept low, but it volatilizes very readily at a high
heat, causing a heavy loss in silver. Such ores have to be
roasted at a very low heat. This subject is exhaustively treated
in another chapter.
In roasting arsenical pyrites, arsenious oxide, sulphurous chlo-
ride, arsenic chloride, and ferric chloride are formed and volatilized.
In the roasted charge we find ferric oxide, ferric sulphate, ferrous
and ferric chlorides, and some ferric arsenate.
Antimony Sulphide. — This mineral we find quite frequently
in complex silver ores, and if it occurs in large quantities the
roasting has to be conducted very carefully and at a very low
heat on account of its great volatility, which can cause a heavy
loss of silver. During oxidizing it changes to oxide of antimony,
of which a large portion is volatilized as such. During chloridiz-
ing antimony trichloride and sulphurous chloride are volatilized.
In the roasted ore we find the antimony as antimony anti-
monate.
8 HYDROMETALLURGY OF SILVER
Quartz. — We find quartz quite frequently as gangue of the ore.
At a proper roasting temperature quartz remains indifferent, but
at a very bright heat it decomposes the salt, forming sodium
silicate and chlorine. There are works in operation in which,
for want of sulphur in the ore, the chlorination of the silver is
produced partly by this reaction and partly by the chloridizing
action of volatilized salt. It requires a high heat and a large
percentage of salt. Quartz is the most desirable gangue in
chloridizing roasting.
Carbonate of Lime (Lime Rock). — This mineral, which occurs
quite frequently as gangue, or part of the gangue, acts as a rule
unfavorably in chloridizing roasting. It takes an active part in
the process. It combines with the sulphuric acid which is pro-
duced by the combustion of the metal sulphides, and it decom-
poses also the metal sulphates, forming sulphate of lime and
metal oxides, thus preventing them from acting on the salt. It
decomposes also metal chlorides, forming calcium chloride and
metal oxides. Calcium sulphate is indifferent and does not act
on the salt. If there is more carbonate of lime in the ore than
can be converted into sulphate and chloride, part of it will be
found in the roasted ore as caustic lime, which acts decomposingly
on the silver chloride, especially so in the subsequent treatment
for extraction, causing a poor result. If, however, there are
more sulphides in the ore than necessary to convert the lime into
sulphate and chloride, usually a good chlorination of the silver
can be obtained, with the further advantage that the final silver
precipitate will be very rich in silver, almost free from base-
metal sulphides, and easily convertible into metallic silver of great
fineness. Therefore, if lime is present in the ore in moderate
quantities it is beneficial to chloridizing roasting. The loss of
silver by volatilization will be found moderate, as most of the
volatile chlorides are converted by the lime into oxides, which
then are not volatile and will not induce silver chloride to vola-
tilize.
Porphyry, Clay, Slate, and Other Gangue Minerals Containing
Alumina. — F. Sustersic made the very interesting observation
that under certain conditions a great loss of silver may be
caused by the presence of alumina. The chlorine acts on the
alumina, forming aluminum chloride, which is extremely volatile
and induces the silver to volatilize. The conditions under which
THEORY OF CHLORIDIZING ROASTING 9
this unfavorable reaction takes place were not ascertained. As a
rule the gangue minerals named in this paragraph are more or
less indifferent; and do not exercise a bad influence in chloridizing
roasting.
CLASSIFICATION OF ORES IN RELATION TO
CHLORIDIZING ROASTING
By the above-described behavior of the different minerals
in chloridizing roasting it is apparent that chloridizing roasting
of complex silver ores is undoubtedly one of the most delicate of
metallurgical operations. The treatment has to be modified in
accordance with the character of the ore, and the character of an
ore in relation to chloridizing roasting depends on the nature of
the different sulphide minerals and the gangue accompanying
them. The sulphide minerals can be classified as:
(1) Those, like iron and copper pyrites, gray copper ore, silver
copper glance, and argentite, which form in roasting sulphates
which act on the sodium chloride and liberate the chlorine.
(2) Those, like galena and zinc blende, which form sulphates
remaining indifferent to sodium chloride.
(3) Antimonial and arsenical silver minerals, which form
antimonates and arsenates of silver.
The gangue either remains indifferent, like quartz and por-
phyry, or it takes an active part, like limestone, and minerals con-
taining magnesia.
If an ore consists of minerals of the first class, together with
an indifferent gangue, chloridizing roasting offers no difficulties
nor does it require much skill, and a high chlorination can be
obtained without much loss of silver by volatilization; nor does
it matter whether the salt is added to the charge before enter-
ing the furnace or after it has been subjected to a partial oxi-
dizing roasting.
The process of chloridizing roasting becomes more difficult if
one or both minerals of the second class are present in large
quantities, even if associated with an indifferent gangue. The
roasting of this class of ore is elaborately treated in Chapter XI.
With such ores the time the salt is added becomes very impor-
tant. If added before the charge enters the furnace a very
inferior chlorination is obtained, as is also the case if the salt is
added before the oxidizing period has sufficiently advanced, or if
10 HYDROMETALLURGY OF SILVER
it is added when the period has too far advanced. The tempera-
ture and air supply require much attention.
The roasting is not less difficult if all three classes are repre-
sented, especially in connection with a gangue like limestone,
which takes an active and often injurious part in the operation.
This class of ore is treated elaborately in Chapter XII.
II
CRUSHING OF THE ORE
THE fineness to which an ore has to be reduced in order to
give the best roasting result depends on the chemical and physical
character of the material. As a rule, finely pulverized ore roasts
quickly and gives a better result than a coarser material. Ores
which decrepitate when charged in the furnace, or ores which
during the combustion of the sulphur swell and disintegrate like
iron pyrites, can be crushed rather coarse and still will give good
chloridizing results. The ore of Sombrerete, Zacatecas, Mexico,
gave good roasting results if crushed through a screen with 10,
even with 8, meshes to the linear inch, though the ore contained
much zinc blende and galena. The zinc blende, however, was of
that kind which decrepitates, and besides, the ore was crushed in
a stamp battery. In crushing in a battery the larger portion of
the material is much finer than the size of the meshes calls for.
This is particularly the case with heavy ores. It is doubtful if
the same good result could have been obtained if the same ore
had been crushed through rolls, because rolls produce a pulp
much more uniform in size, with a much larger percentage corre-
sponding in size with the size of the meshes of the screen.
I made some experiments in this direction with ore of the
San Francisco del Oro mine, near Santa Barbara and Parral,
Chihuahua, Mexico. The ore consists principally of a very dense
zinc blende and finely divided galena. The zinc blende did not
decrepitate. The zinc blende and the galena were the principal
silver-bearing minerals of the ore.
A series of roasting experiments was made with ore crushed
through 20- and through 40-mesh screens. The ore was crushed
in a stamp battery. It was found that the ore crushed through
20-mesh required a much longer time and was 27 per cent, less
chloridized than the ore crushed through the 40-mesh screen.
The material which passes through a battery screen of certain
11
12
HYDROMETALLURGY OF SILVER
size is much finer than the size of the meshes. Heavy ore makes
a much finer pulp through the same screen than lighter ore.
The pulp of the Del Oro ore, obtained by crushing through
battery screens No. 20 and No. 40, was sifted through sieves of
different fineness, and the following figures obtained:
BATTERY PULP
WHEN SIFTED
THROUGH SIEVE
CRUSHED
THROUGH SCREEN
No. 20
CRUSHED
THROUGH SCREEN
No. 40
CRUSHED
THROUGH SCREEN
No. 20
CRUSHED
THROUGH SCREEN
No. 40
PERCENTAGE OF
MATERIAL PASS-
ING THROUGH
THE SIEVE
PERCENTAGE OF
MATERIAL PASS-
ING THROUGH
THE SIEVE
PERCENTAGE OF
MATERIAL
REMAINING ON
THE SIEVE
PERCENTAGE OF
MATERIAL
REMAINING ON
THE SIEVE
No. 30
No. 40
No. 60
No. 80
No. 90
93.8
87.3
78.8
71.2
67.1
100
100
98.95
93.80
90.50
6.2
12.7
20.2
28.7
32.9
0.0
0.0
1.05
6.20
9.50
These figures show how exceedingly fine a heavy ore is crushed
in a battery, even through a screen with comparatively coarse
meshes. Though 67.1 per cent, of the material which was crushed
through screen No. 20 was finer than sieve No. 90, the average
chlorination of quite a number of compared roastings was 27 per
cent, less than that of ore crushed through battery screen No. 40.
This indicates how essential it is to crush such ores fine.
It is frequently argued in favor of coarse crushing that coarser
crushed ore permits in the subsequent lixiviation a free percola-
tion of the solution.
While to a certain extent coarsely crushed ore permits a
somewhat quicker filtration, the increase (if extremes are avoided)
is slight and of not much practical value. If a finely crushed ore
filters too slow for an extraction by filtration it will filter too
slow if it is crushed coarser, because in crushing always a certain
amount of very fine powder (slime) is formed, no matter what
kind of a pulverizing machine is used, and if the nature of the
ore is such as not to undergo much of a physical change in roast-
ing, the pulp in either case will contain sufficient slimes to inter-
fere with a free percolation. A free percolation does not depend
on the coarseness of the pulp nor on the nature of the gangue;
it depends almost entirely on the nature of the sulphides and on
the proportion of metal sulphides and the gangue. Besides the
chemical changes which an ore undergoes during chloridizing
CRUSHING OF THE ORE 13
roasting, a change of its physical condition also takes place.
Lead sulphate, which is formed in roasting, melts easily at a
roasting temperature and is absorbed by the gangue and metal
oxides. The same is the case with cuprous and cupric and with
ferrous and ferric chlorides. They melt even below red heat and
also penetrate the ore. By doing so, these metal salts collect
all the dusty particles or slimes of the gangue and metal oxides
into small porous globules and flakes, in which changed condi-
tion the ore permits a free percolation. This is the cause why a
chloridized ore filters so much better than a raw ore, and if the
ore contains a sufficient amount of metal sulphides it will filter
well whether crushed very fine or whether it is crushed coarser.
If finely pulverized the conditions for the chemical reactions,
however, are much more favorable.
The melting of the metal chlorides and lead sulphate and
their absorption by the ore causes the loosening and swelling of
the charge, making it what is called " woolly," during the chlorid-
izing period. It assumes a moist appearance and can be stirred
without dusting, and does not evade the hoe as during the oxi-
dizing period, but can be banked and collected into a pile. To
maintain this condition the charge has to be agitated from time
to time, otherwise a crust will be formed on the surface.
If for want of a sufficient amount of sulphides in the ore the
formed chlorides and sulphates are insufficient to cause this
physical change, the ore will remain dusty, run like water on the
cooling floor, and will filter very slowly. On the other hand, if
the amount of sulphides, especially lead sulphide, is too great
in proportion to the gangue and metal oxides, the latter will
get so saturated that they cannot maintain their loose condition,
and form lumps. If the temperature is kept moderate these
molten chlorides and lead sulphates will act like a cement, but will
not go into chemical combination with the silica, and in most
cases the lumps will be found to be porous and soft and as well
roasted as the finer part. But if the heat is kept too strong
silicates will be formed and the lumps will become dense and
hard, and the chlorine will be unable to penetrate them and act
on the silver. Particles of undecomposed sulphides will be en-
closed in them and cause a poor extraction. The silver can be
extracted from such lumps only if they are ground and reroasted
with steam, by which hydrochloric acid is formed, which acts on
14 HYDROMETALLURGY OF SILVER
the silicates. Without steam only a small percentage of the
silver can be extracted.
Iron sulphides do not participate so much in changing the
physical condition of the ore as lead or copper sulphides do,
because the chlorides of iron easily give off their chlorine and
change into oxide, which then acts like gangue. The most effec-
tive agent is the lead sulphide, the main part of which is changed
into sulphate, which is but very little volatile at a roasting heat
and does not undergo any further changes, thus much improving
the filtering quality of an ore.
In working the refuse dump of the Cusihuiriachic mine, Chihua-
hua, Mexico, containing from 25 to 30 oz. silver per ton, I found
that, while a satisfactory chlorination of the silver could be
obtained, the silver could not profitably be extracted on account
of the exceedingly slow filtration caused by too great an excess of
porphyry gangue. It occurred to me to add a small percentage
of galena, and the effect was very gratifying — the ore filtered
well. Later the slow filtration trouble was overcome by applying
trough lixiviation.
Ill
PERCENTAGE OF SALT REQUIRED
IF all the chlorine of the salt could be transferred to the silver
only an insignificant amount of salt would be required, but as
other metals, which usually are present in much larger quantities
than silver, are also chloridized, a correspondingly large percent-
age of salt has to be added to the ore. The amount to be added
depends on the nature of the ore and has to be ascertained em-
pirically in each individual case. It is best to commence with a
high percentage, say 10 per cent., of salt, and to reduce the salt 1
per cent, in each succeeding roasting charge until 3 per cent, is
reached. Ores which can be chloridized with less than 3 per
cent, are very rare. The roasted charges are tested in the lab-
oratory for silver chloride. It will be found in most cases that
10 per cent, of salt does not produce a higher chlorination than
6 or 5 per cent., and the experimenter will decide on the least
amount of salt which produces as good a chlorination as the next
larger amount, and will adopt that percentage. There are
instances, however, where it will be found of advantage not to
produce the highest possible chlorination, but to be contented
with a somewhat inferior extraction. This is the case when the
cost of the extra amount of salt exceeds the value of the additional
amount of silver gained. This occurs usually in treating the
lower grade ores in remote localities, where the price of salt is
high.
Ores containing a large percentage of lead and zinc require
less salt than ores rich in iron and copper sulphides, because the
main part of the lead is converted into lead sulphate, which re-
mains indifferent during the chloridizing period and does not
consume any chlorine. This is also the case with zinc, which is
mostly converted into zinc sulphate and oxide, which remain
indifferent. Most of the iron and copper, however, is converted
first into chlorides before they change into oxides, and of course
15
16 HYDROMETALLURGY OF SILVER
these are heavy consumers of chlorine, and the ore therefore re-
quires more salt. For instance, the ore of the San Francisco del
Oro mine in Mexico, which is very heavily mineralized, containing
zinc 24.08 per cent., lead 11.92 per cent., iron 7 per cent., copper
0.5 per cent., and sulphur 21.35 per cent., required only 3J to 4
per cent, of salt.
An excess of salt does not improve chlorination; on the con-
trary, in many instances I have observed that the chlorination
already gained was reduced by adding more salt. For this and
for economical reasons an excess, therefore, should be avoided,
especially as undecomposed salt in the roasted ore is not advan-
tageous in the subsequent extraction.
THE PROPER TIME TO ADD THE SALT
The generation of chlorine in the furnace does not commence
until, by the oxidation of the sulphur, metal sulphates have formed,
which then act on the salt. The first part of the process, there-
fore, is an oxidizing process, whether the ore contains salt or not,
and in this respect it would be immaterial at what time the salt
were added. That the ore sustains a heavier loss of silver by vola-
tilization if the salt is added before the oxidizing period is not
conclusively proved, and actually there is no reason for it.
Chlorides are not formed until the sulphates are formed, and
therefore the presence of salt cannot cause a greater volatiliza-
tion of the silver. An ore which is apt to lose silver on account
of its arsenic and antimony sustains the larger part of its loss
during oxidizing roasting.
There are ores, however, which cannot be chloridized success-
fully if the salt is added to the ore in the beginning. This is the
case with ores which contain a large percentage of a dense argen-
tiferous zinc blende, or argentiferous galena, or both, as the prin-
cipal silver-bearing minerals of the ore. The reason why such
ores have to be first subjected to an oxidizing roasting before the
salt is added is the following:
Zinc blende, if subjected to oxidizing roasting, changes into
zinc oxide and zinc sulphate, while sulphurous acid escapes.
The process of oxidizing the zinc blende progresses but slowly,
especially if the mineral is very dense. Iron and copper sulphides,
on the other hand, oxidize easily and are converted into sulphates
PERCENTAGE OF SALT REQUIRED 17
long before this is the case with the zinc and lead sulphides.
Zinc and lead sulphates do not act decomposingly on the salt,
while iron sulphate does so energetically. Now, if the ore and
salt are charged together we will find that the iron sulphate, as
soon as it is formed, will act on the salt, producing chlorine and
transforming itself into chloride and oxide. The chlorides of
iron are volatile, and also give off the chlorine, changing into oxide.
While this process is going on the zinc and lead sulphides are only
partly oxidized, and as the chlorine in roasting has but very little
effect on the raw zinc blende and galena, the silver contained
therein will not be chloridized by the time the generation of
chlorine and the action of iron chloride has ceased. The conse-
quence is a very inferior roasting result. If, however, the ore is
charged into the furnace without salt and subjected to an oxi-
dizing roasting until the zinc and lead sulphides are oxidized, • or
to a certain extent oxidized, and then the salt is added, the gen-
erated chlorine and the iron chlorides will find the silver in a state
in which it will combine with the chlorine. Iron sulphate re-
quires considerable heat to be decomposed directly into oxide and
sulphuric acid, and if the heat during oxidizing roasting is kept
low, there will be sufficient iron sulphate in the charge to decom-
pose the salt, and a quite satisfactory chlorination of the silver
will be effected. To this class of ores belong those of the San
Francisco del Oro mine, Chihuahua, Mexico, and of Sombrerete,
Zacatecas, Mexico.
Another instance of great difference in the behavior of the ore,
whether the salt was added to the ore in the battery or during
the oxidizing period in the furnace, I experienced during my
investigation of the chloridizing roasting of the calcareous arseni-
cal silver ore at Yedras, Sinaloa, Mexico. The gangue of this
ore consisted of silicious limestone and calcspar, while the ore
proper consisted of argentiferous arsenical pyrites, a moderate
amount of fine-grained black zinc blende, arsenical fahlerz, and
some iron pyrites. When the roasting was done in the Bruckner
furnace there was a marked difference in the behavior of the ore.
When the salt was added in the battery, the ore swelled, became
woolly, kept on one side of the revolving furnace, and when
discharged did not dust and remained in a pile on the cooling
floor. When the salt was added toward the end of the oxidizing
period, the ore did not assume the moist appearance so charac-
18 HYDROMETALLURGY OF SILVER
teristic in chloridizing roasting, but remained very loose and
level in the revolving cylinder, and when discharged made much
dust and ran on the cooling floor like water. The percentage of
silver chlorination was in both cases about the same, but the ore
which contained the salt at the beginning formed a very large
amount of hard balls, which increased in size as the roasting
progressed. They consisted of concentric layers and were smooth
and hard. They were well chloridized, but the silver could not
be extracted unless they were first pulverized, as they were too
dense to permit the solution to percolate through them.
The sulphureted part of the ore had no tendency to form
lumps, as numerous experiments with concentrates of the same
ore showed. In this case we have an instance in which the time
of adding the salt was conditioned by the nature of the gangue
(see Chapter XII).
If the salt is added later, it is not necessary to dry and pul-
verize it; in fact it is better not to do it. It saves expense,
and, besides, it is difficult to spread it uniformly over the charge,
and in places where more salt drops it is apt to form lumps. The
action of finely pulverized salt commences immediately on touch-
ing the ore, and in doing so it becomes sticky, which makes it
difficult to divide and to mix it evenly. This is still more the
case in a Bruckner furnace. If coarse salt is added, the crystals,
which usually are of the size of beans and have more or less moist-
ure, coming in contact with the hot ore decrepitate quite rapidly.
The particles fly in all directions, striking the roof and sides and
falling back to the ore. When decrepitation ceases the salt
will be found much more evenly scattered over the charge than
this can be done by a shovel, and the disintegrated particles are
small enough for the purpose. The chemical action does not
commence quite as soon as with pulverized salt, and a much
better mixing can be secured. Of course, larger lumps of crystals
cemented together, or pieces of salt crust, have to be mashed
first. Salt fuses and is absorbed by the ore, thus coming in inti-
mate contact with the sulphates. A Bruckner furnace should
not be set to revolve until decrepitation ceases.
To extend the oxidizing roasting to such a degree as to produce
a "dead roast," that is, to convert all the convertible sulphates
into oxides and then to produce the chlorination by an addition
of a mixture of calcined copperas (ferrous sulphate) and salt, is
PERCENTAGE OF SALT REQUIRED 19
by far too slow and expensive a method to be adopted in prac-
tice.
To add the salt during crushing produces a very uniform
mixture of ore and salt and simplifies operations in roasting, for
which reason it is preferable to do so if the nature of the ore
permits it; still it is not frequently done in practice, unless the
construction of the furnace demands it, because by crushing ore
and salt together the crushing capacity of the machinery is
reduced by the amount of salt added, and even more if the latter
is not previously very well dried.
IV
LOSS OF SILVER BY VOLATILIZATION
SILVER chloride as such is not volatile, but if influenced by
the volatilization of other chlorides it becomes volatile. A high
heat, therefore, indirectly causes a larger loss of silver by the
expulsion of larger quantities of volatile chlorides. Other condi-
tions being equal we shall always find the loss of silver to be in
direct proportion to the chemical loss in weight an ore sustains.
In other words, the charge of the same ore that during roasting
sustains the least chemical loss in weight sustains also the least
loss of silver by volatilization. The term "chemical loss in
weight" is used in distinction to the loss an ore sustains during
roasting by dusting, which is a mechanical loss.
The logical consequence of the above facts is that the operator,
while he endeavors to obtain a high silver chlorination, should be
at the same time careful to expel as little as possible of the volatile
chlorides. He will be greatly assisted in this endeavor by keeping
the ore in a thick layer, and by using low heat and plenty of air.
If a small charge is thinly spread over a large hearth more volatile
chlorides will be expelled, and the ore will lose more in weight
and in silver than when a larger charge is roasted in the same
furnace. This is the reason why, as a rule, the loss in weight
and in silver in a large Bruckner furnace, in which the ore lies
two feet thick, is less than in a reverberatory, and why small
samples roasted on a roasting dish in the muffle show so much
greater loss of silver than the same ore does when roasted on a
large scale in the furnace.
It will be found that ore roasted at a low heat with sufficient
air will lose less in weight, because a large part of the volatile
chlorides, which at a higher heat would be expelled, will then
remain in the ore. For amalgamation it is desirable, in fact
necessary, to expel the volatile chlorides as much as possible,
because they take an active part in amalgamation and make the
20
LOSS OF SILVER BY VOLATILIZATION 21
quicksilver smeary and inactive, causing a poor silver extraction
and a very low-grade bullion. These chlorides, however, do not
seriously interfere in the lixiviation process; in fact, it is one of
the principal advantages of lixiviation over amalgamation that
in it the volatile chlorides do not need to be expelled, and therefore
the roasting of most ores, even those rich in arsenic and antimony,
can be conducted with a very small loss of silver by volatilization.
In metallurgical books we always find the great loss of silver
pointed out as an objection to all processes which require chlori-
dizing roasting of the ore. Formerly chloridizing roasting was
principally used and studied in relation to amalgamation; little
or no attention was paid to roasting for lixiviation, or to the fact
that this process allowed a modification of roasting by which its
objectionable features could be obviated.
I made chloridizing roasting the subject of special study, and
found that it could be conducted with just as little loss of silver
as oxidizing roasting, if care was taken to expel as little as pos-
sible of the volatile chlorides. The chemical reaction between
salt and the sulphates takes place at a very low heat, in fact at
a lower heat than is generally believed, while on the other hand
it takes quite a high heat to expel thoroughly the volatile chlo-
rides; therefore, in roasting for lixiviation the temperature can be
kept as low as the nature of the ore permits during oxidizing, and
lower still during chloridizing, and yet have the ore well prepared
for the subsequent extraction of the silver. During chloridizing
the ore ought to be kept in a thick layer and stirred only at
intervals to diminish the volatilization of the chlorides.
In the; old method of chloridizing roasting the aim was to free
the ore by heat from metal chlorides that are objectionable for the
subsequent extraction of the silver, while in the new method the
aim is to retain in the ore as much of the chlorides as possible
and to remove them by leaching with water previous to the ex-
traction of the silver. If we take into consideration the fact that
the otherwise not volatile silver chloride becomes volatile by the
volatilization of other metal chlorides, it is quite logical that the
volatilization of the silver will be greatly reduced by the modified
method.
In roasting the calcareous arsenical silver ore at Yedras,
Sinaloa, Mexico, by the modified method, with plenty of air, the
loss in weight was only 3.5 per cent, and the loss in silver by
22 HYDROMETALLURGY OF SILVER
volatilization only 1.8 per cent., while if roasted by the old method
the loss in weight was found to be from 7 to 13 per cent., while the
loss of silver was 15 to 25 per cent, and more.
METHOD OF ASCERTAINING THE Loss OF SILVER BY
VOLATILIZATION
In order to roast skilfully it is of great importance to ascer-
tain frequently the loss of silver by volatilization, but to do
this it is necessary to know the loss in weight the ore sustains.
This, however, is accompanied with great difficulty if it is done
in the old way by actual weighing of the charge before and after
roasting, necessitating the careful cleaning of the furnace and
the dust-chambers before and after the process. In many cases
this is not possible without seriously interfering with the regular
work, and at all events it is accompanied with so much trouble
and expense that if the shrinkage in weight is once ascertained,
this figure is used in all subsequent calculations, though the
conditions under which the roasting is performed, such as heat
and draft or the character of the ore, may have changed. That
such figures are not very reliable will be readily understood, but
still more incorrect is the method some adopt of roasting 10 or
20 grams in the muffle and then taking the difference in weight
before and after roasting as the loss in weight the ore sustains in
roasting; by this means the loss of that particular sample in the
muffle is ascertained, but not the loss the ore would lose in the
furnace. Just as incorrect is the practice of roasting 10 grams in
the muffle, of using the roasted 10 grams for an assay, and of
comparing the assay value per ton with the assay value per ton
of the raw ore. This gives us only the amount of silver this
particular sample lost by volatilization, but it gives no informa-
tion as to how much the ore loses if roasted in the furnace, because
the conditions under which the roasting in the two cases takes
place are very different with regard to temperature, draft, time,
and thickness of the layer.
To conduct the roasting properly it is not of great importance
to know how much the ore loses by dusting, for this is merely a
mechanical loss, and the fine ore particles carried away by the
draft are easily collected in dust-chambers. The loss due to the
volatilization of the chlorides is the serious one. These fumes
LOSS OF SILVER BY VOLATILIZATION 23
are often richer in silver than the ore, are difficult to collect, and
easily escape. We have, therefore, to find how much the ore
in the furnace loses in weight by volatilization in order to ob-
tain a correct basis for a calculation of the loss of silver in roast-
ing.
I adopted the following method, which gives sufficiently correct
results for practical purposes, can be performed in the assay
office in a few hours, and is at all events more correct than if the
loss in weight of the ore is ascertained by actual weighing of the
charge and flue-dust.
Ten grams of the raw pulp, containing the same percentage
of salt as the ore in the furnace, are placed in a roasting dish
and roasted in the muffle for half an hour or an hour; then the
sample is removed from the muffle, allowed to cool, weighed,
returned to the muffle, roasted again for half an hour, and then
weighed again. This is repeated -until two weighings are alike,
or until in the last half-hour the ore does not lose more than
2 or 3 mg.; then the difference between the original weight and
that of the last weighing, expressed in percentage, gives the
highest possible loss the raw ore can suffer.
Ten grams of a sample of roasted ore, corresponding with the
sample of raw pulp, are placed in a roasting dish, and also roasted
in the muffle until two weighings agree, or the difference between
two consecutive weighings is not more than 2 or 3 mg. The dif-
ference between the first weighing (10 grams) and the last, ex-
pressed in percentage, gives the weight which the roasted ore is
still capable of losing if subjected to prolonged roasting. If we
deduct, therefore, the capable loss from the highest possible loss,
we obtain in percentage the loss in weight the ore has suffered
during roasting in the furnace by volatilization.
In the following, the weighings are given of one of the tests
which I made with ore roasted in Bruckner cylinders at Yedras,
Mexico:
RAW ORE, CONTAINING 7 PER CENT. SALT
Original weight 10 grams.
After 1 hour roasting in the muffle 9.35 "
After J hour more roasting in the muffle 9.23 "
After | hour more roasting in the muffle 9.21 "
Ten grams — 9.21 grams = 0.79 grams = 7.9 per cent, highest
possible loss in weight.
24
HYDROMETALLURGY OF SILVER
ROASTED ORE
Original weight •••••• 10 grams.
After 1 hour roasting in the muffle 9.65 "
After ^ hour more roasting in the muffle 9.51 "
After \ hour more roasting in the muffle 9.50 "
Ten grams — 9.5 grams = 0.5 grams = 5 per cent, loss, which
the ore, roasted in the furnace, was still capable of sustaining
by dead roasting.
Highest possible loss of raw ore 7.9 per cent.
Capable loss of roasted ore 5.0 "
Actual loss in weight in the furnace 2.9 per cent.
The gangue of the Yedras ore is limestone. Agreeing weights,
however, are more quickly obtained if ores have quartz gangue,
while ores containing considerable manganese take a longer time,
and require more patience. It is advisable to pulverize the ore
carefully once or twice in a porcelain mortar during the test, in
order to break up small lumps which have formed.
As this test is so quickly and easily done, it gives the metal-
lurgist the means of ascertaining the most favorable tempera-
ture and proper time, and of controlling the work of the man
in charge of the furnace. The mere difference of assay value
between raw and roasted ore is no guide, as can be seen in the
following table, in which the results of a few tests are given,
which I made with the ore of the Cusihuiriachic Silver Mining
Company, Chihuahua, Mexico:
ASSAY
VALUE OF
RAW ORE
PER TON
OUNCES
ASSAY
VALUE OF
ROASTED
ORE
PER TON
OUNCES
HIGHEST
POSSIBLE
Loss OF
RAW ORE
PER CENT.
CAPABLE
Loss OF
ROASTED
ORE
PER CENT.
ACTUAL
Loss
IN WEIGHT
SUSTAINED
IN ROASTING
PER CENT.
Loss OF
SILVER BY
VOLATILI-
ZATION
PER CENT.
REMARKS
46.0
43.6
7.0
6.0
1.0
6.2
41.8
41.6
7.3
5.5
1.8
2.2
43.2
41.0
7.5
5.3
2.2
7.2
Roasted in "C.'s" shift
43.2
41.6
7.5
6.2
1.3
4.9
Roasted in "L.'s" shift
51.0
50.0
7.8
6.6
1.2
3.1
The roasting was done in Howell furnaces. Each of the
above tests was made with average samples of a whole day's
roasting. The third and fourth sample, however, represent the
ore roasted in one day, one roasted in C.'s shift (night) and the
other in L.'s shift (day). Corresponding samples were taken of
the raw ore. The salt was added to the ore in the battery. The
LOSS OF SILVER BY VOLATILIZATION 25
assay showed that the value of the raw ore was in both shifts
the same, while the assay value of the roasted ore of both men
was nearly the same, L.'s assaying only 0.6 oz. more. Judging
by the assays, we are apt to think that the work of both men
was nearly alike, but by referring to the column showing the
percentage of silver volatilization, we find C. lost 7.2 per cent,
while L. only 4.9 per cent, silver. These figures also tell the cause
why C. lost more silver. The ore roasted by him showed a capable
loss in weight of 5.3 per cent, while that roasted by L. showed
a capable loss in weight of 6.2 per cent. C. therefore roasted at
too high a heat, expelling unnecessarily more volatile chlorides,
and by doing so increased the loss of silver.
V
METHODS OF ROASTING
CHLORIDIZING SELF-ROASTING
THIS mode of roasting can only be successfully performed with
highly sulphureted ore and in a furnace the construction of
which permits the roasting of large charges, like the Bruckner
type of furnaces. We have seen above that in roasting for the
process of lixiviation with sodium hyposulphite it is not necessary
to expel the metal chlorides by increasing the heat to bright red
toward the end of the chloridizing period, but that, on the con-
trary, the roasting should be conducted at a low heat to the very
end, to retain in the roasted ore as much of the metal chlorides
as possible in order to reduce the loss of silver by volatilization.
Reflecting on this principle, it occurred to me, while roast-
ing heavily sulphureted ore in a Bruckner furnace, that the
charge if once ignited may, by the oxidation of the sulphides,
produce and keep in store sufficient heat to finish the chloridiz-
ing part of the process without applying any additional heat.
Experiments showed that this could be successfully done, and
that not only was 50 per cent, of the fuel saved, but that, while
the chlorination of the silver was 5 per cent, higher, the loss of
silver by volatilization was materially less than by applying a
second fire. It was possible in this way to roast 10.6 tons of ore
with one cord of wood.
The charges should not be smaller than 4J to 5 tons, otherwise
the heat stored in the ore will die out before the roasting is finished.
When the furnace is charged, a strong fire is kept up until the ore
has fairly started to roast; then the fire is allowed to go out, or
if necessary pulled out, and the fire-door left open to allow a
sufficient supply of air to pass through the furnace. The heat
gradually increases though the fire is out. The charge maintains
nearly a horizontal position. In due time the ore loses its bright-
ness, increases in volume, and begins to assume a more erect
26
METHODS OF ROASTING 27
position, leaning against that side which moves upward. The
chloridizing period has commenced. While during oxidizing the
ore looks bright and the furnace lining dark, just the reverse
can be observed during chloridizing: the surface of the ore looks
dark while the lining, emerging from the ore, looks red. Of
course, that part of the ore which is brought up by the motion
of the furnace is also red, but it quickly darkens.
It will be found that the chlorination of the silver is finished
before the red heat of the charge has entirely died out, and this
is the proper time to discharge the furnace, in the first place to
avoid loss of time, and secondly to avoid dusting. A chloridized
ore when still red does not dust much in discharging, while when
it gets completely dark, but is still hot, it dusts considerably more
than if handled when quite cool.
If the ore is rich in sulphides, the salt can be added, if required,
during the oxidizing period, but this ought to be done quickly in
order not to cool the furnace too much.
I adopted the term "chloridizing self-roasting" for this mode
of roasting because, after the ore is ignited and the fire is removed,
it passes through the oxidizing and through the chloridizing period
without requiring any further attention. One man can attend
to quite a number of furnaces.
The ore thus roasted is roasted at the lowest possible temper-
ature.
CHLORIDIZING HEAP-ROASTING
If silver ore which has been subjected to chloridizing roast-
ing is left in a pile when discharged from the furnace, it will
retain a dark -red heat for many hours, during which time the
process of chlorination continues. I found that, if the chlorina-
tion of the silver is accomplished in the furnace up to 85 or
90 per cent., the increase in chlorination amounts to respectively
2 and 1 per cent., and that this increase takes place principally
during the first two or three hours. By extending the time only
an insignificant increase takes place. This, however, is different
if the chlorination in the furnace be less advanced at the time
of discharge. In such a case a large increase in chlorination
takes place on the cooling floor. C. A. Stetefeldt made the inter-
esting and valuable observation that even in a very poorly roasted
ore the chlorination of the silver can be brought up to a high
28 HYDROMETALLURGY OF SILVER
percentage if the ore is left in a pile on the cooling floor. In
roasting the ore of the Lexington mine he found it to be of
such a nature that a silver chlorination of only about 47 per
cent, could be obtained in the shaft of the Stetefeldt furnace.
This partially roasted ore was piled -on the cooling floor while his
roasting experiments were going on. An examination of the
roasted ore after twelve hours, however, showed an increase in
chlorination of from 47 to 90 per cent. The ore was too heavily
charged with sulphide to be suitable for a complete roasting in
this furnace, and only a partial oxidation took place, but when
piled in a heap the oxidation continued, forming sulphates which,
acting on the salt, produced the chlorination. The temperature
produced by the slow oxidation was sufficiently high for the
chemical reaction. This observation may lead to the adoption
in practice of a new method of chloridizing roasting, which we
properly may call "chloridizing heap-roasting."
It is apparent that, if a chloridizing roasting could be per-
formed just by exposing the ore to a short roasting in the furnace,
and then leaving it to itself in a pile outside the furnace until
cool, the advantages gained would be great, metallurgically as
well as financially. This method, however, is only applicable to
ores not too heavily charged with zinc blende and galena, as I
once had the opportunity to convince myself. When experi-
menting with the heavy zinc-lead ore of the San Francisco del Oro
mine, Chihuahua, Mexico, containing 24 to 25 per cent, zinc,
11.9 per cent, lead, 7 iron, and 21 sulphur, I also tried the Stete-
feldt furnace. This was done more to obtain positive figures and
a complete record of my investigation than in expectation of
obtaining satisfactory results. The ore when removed from the
shaft of the furnace emitted large volumes of sulphurous acid gas.
No chlorine could be detected and the chlorination obtained did
not exceed 15 to 16 per cent. The ore was piled on the cooling
floor. There it continued to roast, emitting sulphurous acid fumes
for several days, until it finally cooled without showing a per-
ceptible increase in chlorination.
Heap-roasting was tried again by me, with the ore of Som-
brerete, Zacatecas, Mexico, which contained 8.9 per cent, zinc,
9.5 per cent, lead, 16.8 per cent, iron, and 26.4 per cent, sulphur.
Though this ore is considerably lighter in zinc and lead than the
ore of the San Francisco del Oro mine, it was still too heavy to
METHODS OF ROASTING 29
be tried with the means of a Stetefeldt furnace. In order to
promise success the oxidation of the metal sulphides had to be
brought to a more advanced state than can be done in a Stetefeldt
furnace, especially as the Sombrerete ore also required an oxidiz-
ing roasting to a certain state before adding the salt. The
experiment was made with the aid of reverberatory furnaces.
In three adjoining reverberatory furnaces three charges of one
ton each were oxidized until the color of the ore commenced to
change to brown, but still contained many black particles, and
still smelled quite strongly of sulphur. Then 6 per cent, of salt
was scattered over the surface of the ore. Immediately after
adding the salt the three furnaces were discharged simultaneously,
and the hot ore of the three charges was piled into one heap in
the yard outside the building and left there to chloridize. After
lying for fourteen hours, it was found, by inserting the sampling
rod, that the ore inside the pile was still red hot, and that the
fumes of the sample still smelled strongly of sulphurous acid.
The color had, to a great extent, changed from brown to red.
A test of the sample showed that only 12.6 per cent, of the silver
was chloridized. After twenty-three hours it was found that the
temperature inside the heap was considerably lower, but still
high enough for the generation of chlorine. A distinct odor of
chlorine was emitted from the sample, but none of sulphur. The
chlorination of the silver had increased to 74.2 per cent. After
thirty-eight hours the ore had cooled down to an extent that no
more chemical reaction could take place. The heap was spread
out and sampled. The color of the ore was as red as that of
charges finished in the furnace. The chlorination of the silver
was found to have increased to 85 per cent. There is no reason
why the chlorination could not have been raised to 90 or 95 per
cent, and higher, if the proper temperature could have been
maintained longer, but the heap being so small, containing only
three tons, it lost its heat before the chlorination was finished.
Only during the time of dumping the hot ore on a pile could
the fumes be seen. As soon as the pile was completed visible
fumes ceased to emanate. A strong odor of sulphurous acid
could be observed for quite a number of hours, indicating that
oxidation was still continuing, but no fumes could be seen.
When the chloridizing period commenced the odor of sulphurous
acid ceased, but no odor of chlorine could be noticed in its place,
30 HYDROMETALLURGY OF SILVER
nor did any visible fumes emanate from the pile. But from
a sample taken from the inside of the pile light fumes could be
observed, accompanied by an odor of chlorine, indicating that
no volatile chlorides were emitted from the pile, and that the
generated chlorine went into combination with the metals of the
ore. The ore being undisturbed and in a thick layer, an excellent
opportunity existed for this chemical reaction. In roasting in a
reverberatory furnace it can plainly be observed that the ore
on the hearths, even on the chloridizing hearth, will not emit
much visible fume, but as soon as the ore is disturbed by the
movements of the rake heavy fumes will be emitted. Now.
these emanating volatile fumes are the very cause of the volatili-
zation of the silver. It is therefore apparent that, if the crea-
tion of such volatile metal chlorides can be avoided, the loss of
silver will be reduced to the minimum — that is, to the loss which
will occur during oxidizing roasting, and which, in most cases,
is very small.
On examination of the ore roasted in that experimental heap,
it was found that much more metal subchlorides than chlorides
were formed as compared with roasting in the furnace. As a
much better utilization of the chlorine takes place if the ore is
in a heap and left undisturbed than when spread over a hearth
in a comparatively thin layer, it is to be expected that roast-
ing in heaps will require less salt. This agrees with observations
I have made by roasting the same ore in a Bruckner and in a
reverberatory furnace. In the Bruckner furnace less salt was
required while a higher chlorination was obtained, together
with a smaller loss of silver by volatilization. In the Bruckner
furnace the ore is in a much thicker layer than in the reverbera-
tory, which causes the better results.
The experiment at Sombrerete was made under very un-
favorable conditions. The heap was too small, containing only
three tons, and was exposed from all sides to the cooling action
of the air, so that the chemical reaction ceased before the chlori-
nation was completed. Notwithstanding this, the results ob-
tained showed that 85 per cent, of the silver was chloridized, and
if we take into consideration the increased furnace capacity, the
reduction in the consumption of fuel and salt, this method surely
offers sufficient advantages to justify further investigations and
experiments. To maintain favorable conditions the hot ore
METHODS OF ROASTING 31
should be dumped into bins made of bricks or stone masonry,
holding 30 to 40 tons. The number of these bins will depend on
the roasting capacity of the works and on the time a heap will
require to complete the roasting. The bins which have to be
placed on the cooling floor should be open on the side toward
the cooling floor, or provided with a good sized iron door, to per-
mit free access, because chloridized ore as a rule does not run
and has to be poked down. The top of the bins should be pro-
vided with hoods, to take off the sulphur gas which will emanate
from the ore for some time.
Chloridizing heap-roasting may prove to be the most rational
mode of chloridizing roasting. A higher percentage of silver
will be chloridized with less loss of silver and at a smaller cost
than if the roasting is finished in the furnace, no matter what
type of roasting furnace may be used. Of course, the ore has to
contain sufficient sulphur — not less than 8 to 10 per cent.
The reverberatory furnaces at Sombrerete roasted from 60 to
80 tons of ore per day, and the space available as a cooling floor
was inconveniently small for the regular work, and a repetition
of the experiment in a proper kiln was, therefore, not practicable.
CHLORIDIZING ROASTING WITH STEAM
If steam is admitted into the furnace during the chloridizing
period, it forms hydrochloric acid, which decomposes the sul-
phides, expels arsenic and antimony, and chloridizes the silver
with great energy, even metallic silver, on which chlorine acts
but slowly. It acts also on metal silicates and chloridizes the
silver contained therein, which otherwise would remain entirely
indifferent to the action of the chlorine. The heavy zinc-lead ore
of the San Francisco del Oro mine, Chihuahua, Mexico, when
in course of experiments it was passed through the Stetefeldt
furnace, showed a chlorination of only 15 to 16 per cent.
It still contained 8 per cent, sulphur, and in order to bring the ore
in better condition for the extraction of the silver it was re-
roasted. The chlorination, however, could only be increased to
about 44.2 per cent. On examination of the ore as it came from
the Stetefeldt furnace it was found that by dropping through the
shaft the main portion of the ore was transformed into minute
globules, which showed that the ore was partially slagged, and
32 HYDROMETALLURGY OF SILVER
the silver contained in these globules resisted the action of the
chlorine. After reroasting, these globules felt between the
fingers just as sharp and glassy as before, but when the reroasting
was done in presence of steam the result was different. The
chlorination increased from 15 to 66.6 per cent, and the globules
became soft and could be powdered between the fingers. Of
course, a chlorination of 66.6 per cent, is very inferior, but this
fact does not interest us just now. The ore had been spoiled in
the Stetefeldt furnace, which made it impossible to produce a
satisfactory chlorination. The present interest is the fact that
these experiments demonstrated the beneficial effect of steam
in roasting. Without steam the chlorination was only 44.2 per
cent, while with steam it was 66.2 per cent. These figures repre-
sent the average of a considerable number of charges. The
globules which remained unchanged when roasted without steam
became soft and assumed the color of roasted ore.
The same conditions were maintained in both cases, with
regard to the percentage of salt to the temperature applied, etc.
The improved results can therefore be credited solely to the
action of the steam .
If an ore is rich in lead it is hardly possible to avoid the for-
mation of lead silicate during the oxidizing period, and the silver
contained therein will not be chloridized during the subsequent
chloridizing period, and consequently will enrich the residues
and be lost. Roasting with steam is, therefore, much to be
recommended for ores containing galena, especially if the galena
is rich in silver, which is very often the case in complex ores.
Objections have been frequently made against roasting with
steam, based on the assumption that it much increases the consump-
tion of fuel, but in actual practice it will be found that the increased
consumption is not serious at all. Waste steam from the engine
can be used, but even if live steam is applied it is not necessary
to use it in such volumes as to cause a marked drop in the tem-
perature of the furnace. A moderate application answers the
purpose. Sometimes the mere keeping of water in the ash-pit
has a decidedly beneficial effect. The steam has to enter the
furnace at the fire end and under the flame, so that it comes
well in contact with the ore. The steam becomes superheated
mostly at the expense of the ore next to the fire-bridge, thus
preventing an overheating of that part of the charge.
METHODS OF ROASTING 33
The use of steam may also greatly reduce the loss of silver by
volatilization. This is mostly noticeable with ore containing rich
antimonial fahlerz and zinc blende besides antimonial galena.
In working the ores of the Silver King mine of Arizona, I had
the opportunity to make very interesting observations with regard
to the effect of steam in reducing the loss of silver by volatilization.
As steam has not on all kinds of ore such a striking effect as
in this case, it will be instructive to give a short description of
the Silver King ore. This was a complex ore, and consisted of
the following silver-bearing minerals:
(1) Native silver in close contact with fahlerz, silver copper
glance, zinc blende, and in some instances with galena. It was
brittle enough, so that a large part of it was pulverized in the
battery. It occurred in the shape of wire, flakes, solid grains, and
in large chunks, and in such quantities that it had to be removed
from the mortars of the battery twice a week by means of shovels.
This silver was of a bright white color, 0.975 fine, and did not
contain any gold.
(2) Silver copper glance with 70.3 per cent, silver, 9.8 per cent,
copper, 17.4 per cent, sulphur.
(3) Antimonious fahlerz, containing over 3000 oz. of silver
per ton. This mineral was the most important constituent part
of the ore.
(4) Zinc blende, of which there were three varieties:
(a) Zinc blende found in large and quite transparent crystals
of a lustrous green color. This was the poorest of the silver-
bearing minerals of the ore, but it was highly interesting from its
beauty as a specimen. It contained only 10.2 oz. silver per ton.
(b) Brown zinc blende occurred in solid masses and in large
quantities, frequently permeated with wire silver, and con-
tained 97.7 oz. silver per ton.
(c) Black zinc blende was more scarce, and contained 40.8 oz.
silver per ton.
(5) Galena occurred in two varieties: the fine-grained anti-
monious with 185 oz., and the coarsely crystallized with only
29 oz. silver per ton.
(6) Peacock copper ore, with 450.6 oz. silver per ton.
(7) Copper pyrites.
(8) Iron pyrites.
The gangue consisted of quartz, heavyspar and some por-
34 HYDROMETALLURGY OF SILVER
phyry. The average value of the ore as furnished to the mill
was 161.4 oz. per ton.
This ore was roasted with 10 per cent, of salt, in a large size
revolving furnace of the Bruckner type, but with a modification
specially designed by myself for this ore. On account of the
antimonious fahlerz, the antimonious galena and the heavyspar,
the ore caked very easily. For this reason, and to avoid excessive
loss of silver by the antimony, the ore had to be roasted at a
very moderate heat. The furnaces were 16 ft. long. It was
found that the roasting could not be done properly with a furnace
of common construction, with a fireplace only at one end of the
cylinder, as the ore either did not receive enough heat at the
farther end, or, if it did, it was overheated and caked at the end
nearest to the fire. To overcome this difficulty the cylinder was
provided at each end with a fireplace and flue arrangement (see
Figs. 15, 16, and 17). These two fireplaces were worked alter-
nately. After the ore was charged, the furnace was set in slow
revolving motion, and fire kept up in one of the fireplaces. The
flame traversed the furnace, and smoke and gases escaped
through the flue, in front of the opposite fireplace. After a lapse
of one hour, fire was made up in the other fireplace, the damper
reversed, and flame and gases allowed to pass through the fur-
nace in the opposite direction. The changing of the fire was
kept up during the whole time the charge was in the furnace,
only the intervals were not quite as frequent as in the begin-
ning. This system of double fireplace and flues proved to be
of great advantage in securing a very uniform roasting; the
ore from both ends was chloridized up to the same percentage,
while, when the ore was roasted in a furnace with a fireplace at
one end only, the farther end showed a less chlorination of
from 5 to 10 per cent. Besides, it enabled the operator to
roast at a low and uniform heat.
Notwithstanding the capacity of the furnaces to roast at a
moderate and uniform heat, the loss of silver by volatilization
proved to be exorbitant, being not less than 38 per cent., while at
the same time the chlorination was low on account of the large
percentage of metallic silver in the ore, which was but imperfectly
converted into silver chloride by the chlorine. A jet of steam
was then tried, which was applied right under the flame and
directed toward the side where the ore was. There was a steam
METHODS OF ROASTING 35
jet at each end, but only the one was operated which corresponded
with the end at which was the fire. The roasting results thus
obtained were very satisfactory. An average of many charges
showed that the loss by volatilization was reduced from 38 per
cent, to 2 per cent, while the average chlorination of 67 furnace
charges proved to be 94.4 per cent, and in some cases as high as
96.8 per cent.
This roasting example illustrates that with certain ores the
application of steam is of vital importance. The ores of the
Silver King mines could not have been worked by a hydro-
metallurgical method if the steam had not so greatly reduced
the loss of silver, and increased the percentage of chlorination.
There are some ores which do not need steam, but in most
cases a larger or smaller jet of steam, according to the nature of
the ore, does beneficially assist the chemical reactions.
CHLORIDIZING ROASTING OF SILVER ORES
CONTAINING GOLD
There are two combinations of gold and chlorine: the aurous
and the auric chloride. The latter is soluble in water and is
formed when finely divided gold is brought in contact with
chlorine gas at a common or moderately warm temperature.
At a temperature of 230 deg. C. it changes into aurous chloride,
which, however, on further heating, decomposes into metallic
gold and chlorine. Owing to this property of the chlorino com-
pounds of the gold, neither of them will be formed in the furnace
during chloridizing roasting. The temperature in the furnace
is too high for them to exist, and the gold on discharge of the
furnace will be found in the metallic state. The aurous chloride,
which is not soluble in water but is soluble in a solution of sodium
hyposulphite, resists a much higher temperature than does the
auric chloride, and it will form at a temperature much higher
than the decomposing point of the auric, which temperature,
however, has to be kept below red heat.
Based on this property of the gold chlorides I adopted a modus
operandi by which I was able to extract 75 to 80 and even 90
per cent, of the gold contained in the silver ore simultaneously
with the silver by sodium hyposulphite.
If the ore leaving the furnace is not allowed to cool quickly,
36 HYDROMETALLURGY OF SILVER
but, on the contrary, is made to cool slowly by dumping it to a
large pile and leaving it undisturbed until it is cool, which takes
several days, it will be observed that the generation of chlorine
still continues for a considerable time. The cooling of the heap
begins from the outside and progresses toward the inside, and
the chlorine, which is generated at the inside, in escaping will
meet a layer of ore sufficiently cooled for combination with the
gold contained therein. It will form the aurous chloride, because
the temperature is still too high for the auric chloride to exist.
The formation of aurous chloride will progress toward the inside
in proportion to the cooling of the pile. The cooling does not
need to be continued beyond the time when a sample taken from
the inside does not emit any chlorine.
If no precaution is taken to cool the ore slowly, only a small
percentage of the gold will be converted into aurous chloride,
and the gold extraction, therefore, will be very small. The
beneficial effect of slow cooling on the chlorination of the gold
contained in auriferous silver ore can also be observed in ex-
perimenting on a small scale, which will be illustrated by some
results which I recently obtained in conducting some laboratory
investigations respecting the ore of the Lucky Tiger mine, Sonora,
Mexico. An analysis of the sample showed the ore to consist of:
Iron 2.92 per cent.
Zinc 3.36
Lead 1.15
Copper trace.
Antimony trace.
Sulphur 2.54 per cent.
Silica 89.54
Silver 108.16 ounces per ton.
Gold 0.36
Two lots of 100 grams each were roasted with salt on
roasting dishes in the muffle at a dark-red heat. The amount of
salt as well as the temperature and roasting time were for both
lots exactly the same. When roasting was completed one lot
was withdrawn and allowed to cool at a place away from the
muffle, while the other was placed in a hot roasting-dish and
covered with another hot roasting-dish, then removed from the
muffle, but placed right in front of it. Thus the one lot was
allowed to cool quickly, while the other was made to cool slowly.
The quickly cooled ore showed a gold chlorination of 20.8 per
cent, that is, 20.8 per cent, of it could be extracted with sodium
METHODS OF ROASTING 37
hyposulphite, while the slowly cooled ore showed a gold chlori-
nation of 74.7 per cent. The result of this experiment clearly
demonstrates that the chlorination of the gold takes place out-
side the furnace and is caused by slow and gradual cooling. The
conditions in this experiment were not as favorable as they
would have been on a large scale, because the generation of
chlorine in so small a lot as 100 grams ceases soon, while in a
large pile it continues for many hours.
While this mode of roasting gives very satisfactory results
with silver ores containing 1 oz. of gold per ton or less, it will
not be quite satisfactory for ores richer in gold, in which case
a cyanide solution should be applied after the extraction of the
silver. In the above experiment the residues of the quickly
cooled ore were treated with a cyanide solution, by which the
gold extraction was raised from 20.8 per cent to 86.07 per cent.
CHLORIDIZING ROASTING FOR AMALGAMATION
We have seen that the aim in chloridizing roasting for the
extraction of the silver by lixiviation with sodium hyposulphite
is to convert as much as possible of the silver into silver chloride
and at the same time to expel as little as possible of the volatile
base-metal chlorides, in order to reduce the loss of silver by vola-
tilization to the minimum. This modification in chloridizing roast-
ing was an important step forward in the hydrometallurgy of
silver because thereby its weakest and most objectionable feature,
the volatilization of the silver, was reduced to the minimum.
Chloridizing roasting was first devised, studied, and executed
to meet the requirements of barrel and later of pan amalgamation.
For this process not only as much as possible of the silver has
to be converted into silver chloride, but it is of the greatest im-
portance to expel or to decompose the base-metal chlorides, too,
because if this is not done these chlorides will amalgamate with
the mercury as well as the silver and produce an impracticably
large amount of amalgam and a very low-grade bullion. Besides
this, the mercury containing much of such amalgam has no ten-
dency to unite when the pulp is diluted, but remains in minute
globules, which partly float on the surface of the pulp as a dark
scum and partly are carried off with the residue, thus causing a very
large loss of mercury and, of course, of the silver it contains.
38 HYDROMETALLURGY OF SILVER
The mercury loses much of its decomposing energy on silver
chloride; it becomes "foul/7 which results in very rich residues.
In short, it is absolutely necessary in working complex silver ores
by amalgamation to expel and decompose the base-metal chlorides.
This is done by heat.
The salt is added either together with the ore or after the
oxidation of the sulphides has pretty well advanced. During
the oxidizing period the temperature is kept moderate, partly to
prevent caking but mostly to form as much metal sulphates as
possible, because these sulphates will act on the salt and change
into chlorides and also produce chlorine, and, in presence of steam,
hydrochloric acid. If the ore is rich in sulphurets the heat
created by their oxidation can so increase as to make it necessary
to let the fire go out entirely, but close attention has to be paid
to start the fire again as soon as a pronounced decrease of the
temperature can be observed. If an ore charge is allowed to cool
too much it takes considerable time and fuel to bring the temper-
ature up again to the desired degree. During the oxidizing
period the ore ought to be frequently stirred — continually, if
possible. The beginning of the chloridizing period is indicated
by the swelling of the ore; it becomes "woolly" and emits heavy
white fumes. The heat is then gradually increased until toward
the end the charge assumes a bright cherry red. During roasting
the charge is several times turned, that is, the half toward the
fire-bridge is removed toward the flue, and that from the flue
side is brought toward the fire-bridge. This is done in the
following way: The whole charge is raked together into a long
heap extending from the bridge to the flue end. This ridge of
ore is made nearer to the working doors, to lay bare as much of
the hearth as possible. Now, by means of a shovel the ore from
the fire-bridge is moved and dumped back of the ridge, com-
mencing close at the flue and proceeding toward the fire-bridge
until the middle is reached, then the other part of the ridge is
moved toward the bridge. Then the whole charge is spread
again.
During roasting, samples are taken and tested with regard to
the behavior of the ore toward mercury. If it is found that by
adding some water and some mercury to the ore the bright sur-
face of the latter becomes immediately covered with a black
skin, it is an indication that the roasted ore still contains an
METHODS OF ROASTING 39
injurious amount of base-metal chlorides, and roasting has to be
continued at an increased temperature in order to decompose
them.
The practice of decomposing the base-metal chlorides by heat
and increased roasting time is naturally connected with much loss
of silver by volatilization. The late G. Kiistel proposed a much
more rational means than heat to destroy the base-metal chloride.
In his treatise, "Roasting of Gold and Silver Ores," he says:
"It is very difficult to get rid of all the base chlorides; they
are formed under the action of chlorine and hydrochloric acid.
The most of the metal chlorides are volatile, and a part is carried
off through the chimney. Another part of the chlorides gives
off some of its chlorine, whereby sulphates, undecomposed sul-
phurets, antimonates, and arsenates are chloridized. Chlorides
which are disposed to transfer chlorine to other metals in com-
bination with sulphur or arsenic are: the proto chloride of iron
and of copper, the chlorides of zinc, lead, and cobalt. When in
this way the most of the metals are chloridized, the base metals,
principally iron and copper, are losing their chlorine gradually,
being first converted into sub-chlorides, and then into oxides.
The roasting for this purpose must continue with increased heat,
even when the chlorination of the silver is finished. At an
increased heat, the base metal chlorides lose their chlorine, while
the chloride of silver remains undecomposed, unless a very high
temperature should be applied. This process requires a long
time, consequently also more fuel. The decomposition of these
chlorides is greatly assisted by the use of 5 to 6 per cent, of carbo-
nate of lime in a pulverized condition. Lime does not attack the
chloride of silver, but it is not advisable to take too much of it,
as it would interfere to some degree with the amalgamation.
The pulverized lime-rock must be charged toward the end of
the roasting. First, two per cent, is introduced by means of a
scoop, the whole well mixed, and then examined either with
sulphide of sodium or in the following way:
"A small portion of the roasted ore is taken in a porcelain
cup or glass, and mixed with some water by means of a piece of
iron with a clean metallic surface. If the iron appears coated
red with copper, some more lime must be added. In place of
iron — especially if no copper, but some other base metal is
present — some quicksilver is mixed with the sample. In the
40 HYDROMETALLURGY OF SILVER
presence of base-metal chlorides, the quicksilver is coated imme-
diately with a black skin.
"When endeavoring to expel the base metals by heat, the
loss of silver, in presence of much antimony, lead, and copper,
should be investigated very carefully. Under certain circum-
stances it is not uncommon to find a loss of even 50 per cent, of
the silver, if the chloridizing roasting is carried on at a high heat
for a great length of time. The loss increases with the duration
of roasting and with the degree of temperature. When such ore
is under treatment, it is necessary to take samples during the
roasting, and to examine the same for the amount of chloride of
silver, and also for its loss, and to stop roasting when the highest
percentage of chloride of silver is obtained without reference to
the condition of base metals."
Mr. Kiistel proposed a still more rational method for remov-
ing from the roasted ore the base-metal chlorides which are so
obnoxious to amalgamation. Instead of destroying them by
heat and extended roasting time, he removed all soluble chlorides
and sulphate of zinc by leaching the ore with hot water previous
to amalgamation. The ores at Flint, Idaho, turned out such
base amalgam that further working proved to be ruinous; but
after Mr. Kiistel applied his method the ore became most suit-
able for amalgamation, and very satisfactory results were
obtained.
It is apparent that this method is quite an advance in
amalgamation; not only is the amalgamating energy between
silver and mercury much increased, thus resulting in a better
extraction of the silver, but the volatilization of the silver in
the furnace, and the loss of mercury and silver in the process of
amalgamation, are much reduced.
An important point in this process should not be overlooked,
namely, that chloride of silver, while not soluble in water, is
soluble in a concentrated solution of metal chlorides. The dis-
solving energy of such a solution on silver chloride increases with
its concentration and its temperature. Hot water has to be
used in order to remove the lead chloride, and therefore the
first part of the outflowing solution will be as a rule rather con-
centrated, and at the same time, being warm, will dissolve quite
a noticeable amount of silver chloride. This, however, can be
regained by collecting the solution in large tanks and by diluting
METHODS OF ROASTING 41
the same. If sufficient water is added, all the silver chloride will
precipitate, and if enough time can be given will settle on the
bottom together with lead chloride. If there is copper in the
ore the solution should be made to flow through tanks filled with
scrap iron, by which the copper and silver are saved.
Lead chloride amalgamates; lead sulphate does not; and in
roasting for amalgamation it is therefore of importance to con-
vert as much of the lead as possible into sulphate and as little of
it as possible into chloride.
Lead sulphate, if once formed, remains indifferent and un-
changed during the balance of the process. Mr. Kiistel, in roast-
ing the plumbiferous silver ores for amalgamation at Plomosas,
Mexico, made the observation that if the ore was roasted in a
reverberatory with a roof of only 20 or less inches above the roast-
ing floor the bullion obtained contained 500 to 600 parts of lead
in 1000, while if the same ore was roasted in a reverberatory with
high roof, 27 or 30 inches, the bullion obtained was almost free
of lead. In the first instance much lead chloride and oxichloride
was formed, which amalgamated, while in the second instance
nearly all the lead was transformed into lead sulphate. The
reason of the different results is apparent. In the furnace with
the higher roof there was sufficient space left between the flame,
which travels next to the roof, and the surface of the ore to per-
mit a free access of air, and in presence of ample live air more
sulphuric acid is formed, which acts on lead oxide, oxichloride,
and chloride, converting them into sulphate. If the supply of air
is limited by the low arch, insufficient sulphuric acid is formed
to convert all the lead into sulphate. In support of this theory
speaks the fact that C. A. Stetefeldt, by roasting the ore from
Ontario, Utah, in a Stetefeldt furnace, found all the lead contained
in the ore to be in the state of sulphate.
With regard to the extraction of the silver by sodium hypo-
sulphite it is immaterial whether the lead in the roasted ore is in
the state of sulphate or chloride. If cold water is used for re-
moving the base metal chlorides, which is generally the case,
but a small portion of the lead chloride is removed, and the same
is still contained in the ore when the latter is subjected to leach-
ing with sodium hyposulphite, in which solution it is soluble, as
well as the sulphate, which is not soluble in water.
VI
CONSUMPTION OF FUEL
THE fuel mostly used in roasting is wood, not so much so
because it is the fuel with which it is the easiest to regulate the
temperature, but because the mines and works are usually situated
in more or less remote mining districts, where wood is easier and
cheaper to be procured than other fuel. Bituminous coal and
gas from gas producers answer the purpose perfectly well.
The quantity of fuel required depends mostly on the nature
of the ore, but also on the construction of the furnace. Ores
rich in sulphur, especially if a large part of it is combined with
iron, as in iron pyrites, require the least amount, because by
the combustion of the sulphur much heat is developed in the ore
itself, so that the process needs only to be slightly assisted by
fire. In fact, ores with 20 or 22 per cent, sulphur, if roasted in a
Bruckner furnace, need only to be heated until it is ignited. The
balance of the heat required for the oxidation and chlorination
is then furnished by the ore, the same as in chloridizing self-roast-
ing, by which method 10 or more tons of ore can be roasted with
one cord of wood. A highly sulphureted ore which does not con-
tain too much zinc blende or galena but contains considerable
pyrite could be roasted chloridizingly without the use of any fuel
except what is needed to start the furnace. We have seen above,
in heap-roasting, that only a partial oxidation of the ore in the
furnace is required; and if for this purpose a continually dis-
charging furnace of suitable construction is used, like those of
the McDougal type as employed in sulphuric acid factories, or of
the Howell type (if the latter is lined with bricks the whole
length and only a very slight inclination given, and provided
with a ring-flange at both ends so as to retain a somewhat larger
amount of ore in the furnace in order to produce more heat and
to preserve it better) a continuous chloridizing roasting could be
successfully effected without the use of fuel.
These continually discharging furnaces will have to discharge
42
CONSUMPTION OF FUEL 43
into closed pits, which in starting will have to be first well heated
like the furnaces. From these pits or vaults the ore is then
moved and charged into the roasting bins. The vaults should be
sufficiently large to prevent cooling.
Ores, very poor in sulphur, containing only about 2 per cent.,
require much fuel, and if the same is expensive should not be
roasted in a reverberatory furnace, because the conditions pre-
vailing in a reverberatory furnace as regards utilization of heat
are very unfavorable. Heat .penetrates but very slowly into any
pulverized material, especially if the same is left undisturbed.
If a fresh charge is placed on the hearth of a reverberatory and
spread over it, it will cool the bottom, and if it is not a muffle
furnace, the supply of heat from that source will soon end, and
the heating will be effected only on the surface by the radiating
heat of the flame which travels next to the roof of the furnace;
but heat, as stated, penetrates very slowly into pulverized ma-
terial, and therefore only a small percentage of the heat produced
by the fuel will be effective. To better utilize the heat it is
necessary that continually new particles of the ore be brought to
the surface to be exposed to the radiating heat. This, however,
for obvious reasons, can be done but very imperfectly in a hand-
stirred furnace. As soon as the charge becomes hotter than the
bottom of the furnace the bottom will draw heat from the charge,
thus exerting a cooling action on the latter. To make the furnace
long helps, but not very much; and if the roasting of such dry
ore has to be done in a reverberatory furnace it is more rational
to make the hearth short and to build the furnace in two stories,
in which case the roof of the lower hearth will heat the bottom
of the upper one.
The most effective and at the same time the most fuel-saving
roasting furnace for this class of ores is undoubtedly the Stetefeldt
furnace, in which the flame is made to ascend through a high
shaft, while the ore in a fine shower falls down the shaft and
against the flame. There is no other roasting furnace in which
the heat is utilized to such advantage as in the Stetefeldt ; ten to
twelve tons of such dry ore can be roasted with one cord of wood —
a result which cannot be obtained in a reverberatory furnace.
This peculiarity of the Stetefeldt furnace, which is so advantageous
for dry ore, makes it also unsuitable for ores heavily charged with
sulphurets.
44 HYDROMETALLURGY OF SILVER
If circumstances do not allow the erection of a Stetefeldt
furnace, the next best fuel-economizing roasting furnace for such
ores is the Howell -White.
If the ore permits a closer sorting without causing too much
waste of silver, it will be well to do so in order to raise the sulphur
contents of the ore, because this will not only reduce the consump-
tion of fuel, but also improve the extraction and shorten the
roasting time. In case the ore is not very rich, and concentrates
well, it may be advisable to concentrate part of it and to add the
concentrates to the balance of the ore, or to sort close and to
concentrate the second-class ore, just as, according to circum-
stances, is found to be the most rational.
VII
REVERBERATORY FURNACES WORKED BY HAND
THE reverberatory furnace is a horizontal hearth furnace
provided with a fireplace and grate at one end and a flue at the
opposite end, and with working doors on one or on both of the
two long sides. The hearth is separated from the fireplace by
the fire-bridge. It is the oldest and the most primitive type of
roasting furnace, but notwithstanding this it is the furnace
which can be applied to any kind of ore, except those the nature
of which prohibits chloridizing roasting altogether, like ores con-
taining too much lead. Its construction gives the operator full
control over the process. It offers facilities to maintain any con-
dition the nature of an ore requires; and it is, when substantially
built, very durable, requiring little repair, and makes but little
flue-dust. The reverberatory is, in fact, the ideal furnace for
chloridizing roasting, and would be exclusively used for this
purpose if it were not for the fact that it has to be operated by
hand, which makes the cost of roasting too high in localities
where labor is costly. This was the cause which gave the impulse
in the silver districts of the Pacific coast of the United States to
the invention and construction of quite a variety of mechanical
furnaces, all of which are labor-saving, and if applied to the
proper ore do excellent service. They cannot be used, however,
for so many kinds of ore as the reverberatory. Each of them
has its peculiarities, with which the character of the ore has to
comply, and it is therefore of the greatest importance, if mechani-
cal furnaces are to be erected, that a thorough study of the
nature of the ore should precede the selection of the furnace.
The Single-Hearth Reverberatory. — This is the oldest style of
a roasting furnace. Figs. 1 and 2 represent the vertical and the
horizontal section respectively: a, hearth; s, roof; h, fireplace;
it fire-bridge; e and e', flue; b, bottom discharge opening; d, vault
for placing the wheelbarrow to receive the roasted ore; p, charge
45
46
HYDROMETALLURGY OF SILVER
hopper in the roof of the furnace. (This hopper is provided with
a slide which, when drawn, permits the charge to drop into the
furnace. It has to be large enough to hold a full charge, and
ought to hang on a proper framework, so that its weight does not
rest on the roof of the furnace) ; w, buck stays and anchors.
i a a 4?
FIGS. 1 and 2. — SINGLE-HEARTH REVERBERATORY FURNACE.
The single-hearth furnaces are not in use any more for roasting
ore on a large scale. They are too wasteful with regard to fuel;
the heat of the gases is not utilized. We find them, however,
very useful for experimental work and for burning the precipitate
in lixiviation works.
The Two-Story Single-Hearth Furnace. — This is a consider-
able improvement on the single-hearth furnace. It is shown in
REVERBERATORY FURNACES WORKED BY HAND
47
Fig. 3, which represents in a longitudinal section the general
arrangement: a, a are the lower and upper hearth ;r, lower fireplace;
6, flue connecting the lower with the upper hearth; &', flue in the
arch of the upper hearth, whence the gases are led to the dust-
chambers; /, working doors of the upper hearth (the working
doors of the lower hearth are on the opposite side); r', an aux-
iliary fireplace for the upper hearth, which is smaller and is
used only when a fresh charge enters the furnace, to assist in
heating and to ignite the same more quickly.
FIG. 3. — TWO-STORY, SINGLE-HEARTH REVERBERATORY
FURNACE.
As soon as the charge of the lower hearth is finished and re-
moved the upper charge is dumped down through the flue 6. To
facilitate this operation, and to permit the use of a hoe instead of
the slower- working shovel, there is at the end side of the furnace
the door /, through which is inserted the hoe with which the charge
is pushed to drop through b.
The Long Reverberatory Furnace. — A further step in the develop-
ment of the reverberatory furnace was the construction of long
furnaces with three to five hearths on the same level, or in flat
steps of three to five inches rise. The length depends on the nature
of the ore. For highly sulphureted ore, especially if it contains
much iron pyrites, the length may be extended to 50 ft. without
the aid of an auxiliary fireplace. As a rule each hearth is made
10 ft. long and 10 ft. wide. The arch, which can be made rather
high at the fire end, ought to slope down toward the flue end to
48 HYDROMETALLURGY OF SILVER
throw the heat of the moving gases more toward the bottom at
the part of the furnace remote from the fire. If the arch is made
straight, each succeeding hearth should be a few inches above the
preceding, by which the same object is attained. These steps
serve at the same time as a mark for each hearth, and assist in
preventing the charges from getting mixed.
These long furnaces are either built singly or in pairs back
to back, as shown by Figs. 4, 5 and 6. Single furnaces are to
be preferred, as they offer the opportunity of providing work-
ing doors on both sides, which not only facilitates the working
of the charge but also permits a free access of air from both sides,
which is of great advantage. The construction of these single
furnaces, however, is much more costly, and at the same time
requires a good deal more space and consequently much larger
buildings. A reverberatory furnace with working doors on one
side only has two dead places, that is, places which are not reached
by live air, for which reason the process of roasting on these
places is not only retarded, but also such places become, as a
rule, overheated and often cause there caking of the ore. These
places are: the part of the hearth next to the fire-bridge extend-
ing almost to the first working door, and the part along the back
wall of the furnace extending from the fire-bridge to the flue, so
that, if no provision is made for air to enter at these points, which
we very often find to be the case, we have to consider the furnace
to be of faulty construction. The oxygen of the air is the very
life of the process, and if the same is withheld, or its access ob-
structed, the bad effect will invariably manifest itself by an in-
ferior roasting result and a higher loss of silver by volatilization.
These dead spaces are avoided by constructing air channels
leading from the front under the hearth and entering the furnace
through openings in the back wall as shown by b and b' ', Figs. 4
and 6. The fire-bridge is also provided with an air channel, /,
which is in communication with three openings, /', through which
the air enters the furnace. The effect of this additional air supply
is very noticeable. The temperature through the whole width
of each hearth is quite uniform and no overheating next to the
fire-bridge takes place. At the same time the fire-bridge is much
protected by the cooling effect of the air. If steam is to be used
in roasting, the channel / serves for this purpose. The steam
pipe, which is provided with three holes while the end is closed
REVERBERATORY FURNACES WORKED BY HAND
49
—44
50
HYDROMETALLURGY OF SILVER
by a cap, is inserted through /. The holes in the pipe are so
divided that each one is located right under each upraise of the
channel.
The long furnace, as represented by Figs. 4, 5 and 6, was
designed by me for chloridizing the calcareous and arsenical ores
of the Anglo-Mexican Mining Company at Yedras, Sinaloa,
Mexico. This ore was of a very peculiar character, and apt to
sustain an unusually heavy loss of silver if certain conditions in
the treatment were not scrupulously maintained. The dimen-
sions of different parts of this furnace were designed to conform
with the peculiarities of this ore, but the general arrangement
was not altered; and the diagrams will well serve to illustrate the
construction of the long reverberatory furnace.
Section J.K.
FIG. 6. — LONG REVERBERATORY
FURNACE.
It is always advisable to make the fireplace (see K, Fig. 5)
sufficiently wide, because it gives the means to regulate the tem-
perature, and if too narrow may make it impossible to supply
sufficient heat. The heat of the first and second hearths is mostly
supplied by the combustion of the sulphides on the preceding
hearths. Wood requires a wider fireplace than coal; 2 ft. 6 in.
will be sufficient for wood fire. The sides and roof of the fire-
box should be lined with fire-bricks in order to resist better the
heat and the wear caused by wood and tools. The brickwork
encasing the fire-box should be substantial and well braced.
The depth should not be made much longer than the length of
the wood, which is usually cut 4 ft., so that a depth of 5 ft. is
sufficient. It is well to keep water in the ash pit, not only to
lengthen the life of the grate bars but also to make a limited
amount of steam, which benefits the roasting very much.
The top of the fire-bridge (see m, Fig. 5) should not be
too high above the grate, in order that the furnace may receive
REVERBERATORY FURNACES WORKED BY HAND 51
as much as possible of the radiating heat of the fire; the width
differs according to circumstances, and is made from 12 to 18 in.
A 12-in. bridge should be made entirely of fire-bricks, but if wider
can be lined on both sides with them. The space above the
fire-bridge is to be made large enough so that the hot gases can
enter the furnace freely without receiving any back pressure,
which manifests itself by flames and smoke coming out between
the frame and door of the fire-door after each new addition of
wood. In case the space is too small the flame recoils, becomes
short and overheats the fire-box without furnishing sufficient
heat to the parts of the furnace farther off from the fire. If this
space is large enough the flame rolls slowly, touching the roof,
and after a fresh addition of wood extends 25 to 30 ft. into the
furnace. The fire-bridge ought never to be built without air
channels, as described above.
The length of the hearth (see n, Fig. 5) depends, as above
mentioned, entirely on the nature of the ore. If an ore is rich in
sulphur the hearth can be made 50 ft. long, but this is the limit.
It is the heat created by the combustion of the sulphides which
makes the working of such long furnaces possible. An ore poor
in sulphur never could be heated sufficiently to commence roasting
if 40 or 50 ft. away from the fire, .and, therefore, a large part of
the furnace would be inactive, causing only unnecessary extra
labor to move the ore. To insert additional fireplaces does not
offer any advantages; on the contrary, it hinders the execution
of a delicate chloridizing roasting. A very uneven heating of
the charge is caused by them. Near the inserted fireplace the
charge gets hot, and often hotter than it ought at that stage of
roasting, and when moved to the next hearth gets cooler again,
which is not advantageous to chloridizing roasting. The insertion
of additional fireplaces is only justified in mechanical continually
discharging hearth furnaces like the O'Harra and the Ropp fur-
naces, which are made 100 ft. long and even longer, and are very
effective in labor saving and under proper conditions do very
good work; but their applicability is confined, like that of other
continually discharging mechanical furnaces, to certain kinds of
ore, and they do not permit a really delicate roasting. On the
first, or charge hearth, the ore ought to become well heated, so
that, shortly after transferring it to the second hearth, blue
flames can be observed when the ore is stirred. Before it is
52 HYDROMETALLURGY OF SILVER
removed from here the oxidation of the sulphides ought to be
well started.
Of course, it takes experience and skill to judge the proper
length to be given to the furnace for a certain ore. Ores con-
taining 20 to 22 per cent, sulphur, with considerable iron pyrites,
will roast well in a furnace 40 to 50 ft. long, provided the ore does
not contain an excess of zinc blende. For ores with about 8 per
cent, sulphur a furnace 30 ft. long is sufficient, and ores containing
only 2 to 3 per cent, sulphur should not be roasted in a long rever-
beratory, but either in a Stetefeldt or a Howell furnace.
Each hearth of the long reverberatory furnace is made 10 ft.
long. If a step is given to each hearth, it serves as a mark,
but if the whole hearth is level, the points, o, of the pillars of the
front wall mark the lines. It is not advisable to make the hearth
too wide, thus trying to increase the capacity of a furnace. It
should be borne in mind that the charges have to be worked and
moved by hand labor, and that extra long tools are very hard to
handle; they tire out the man, in consequence of which the part
of the charge next to the back wall will be worked less than the
part from the middle toward the front. For the same reason
the handle of the hoe and the shovel should be made of gas pipe,
in order to make them as light as possible. To the end of the
pipe is forged a solid rod 24 to 30 in. long, to which the hoe or
shovel is attached. Nothing smaller than a IJ-in. pipe should
be taken, in order to make the handle sufficiently stiff and at
the same time easy for the workman. The hand has an easier
hold on a IJ-in. handle than on an inch handle.
To make the width of the hearth, measured from the front of
the working door to the back wall, 10 to 11 ft. will be found
convenient, but it should never exceed that. The hight of the
hearth above the working floor should be 2 ft. 9 in. to 3 ft., so
that the workman can throw his weight on the handle of the tool
when required, which assists him much.
Of great importance is it to prepare the bottom well so that
it remains level and does not sink in different places. It would
be too expensive to build the whole part below the hearth solid
with bricks, for which reason only the sides and ends are built
solid, while the inside is filled with stones, gravel and sand.
The filling has to be done carefully, so that no hollow spaces are
left. The filling is commenced with coarse rock of 5 to 6 in.,
REVERBERATORY FURNACES WORKED BY HAND 53
then, after about 8 in. in depth are filled, finer material is used
and washed down between the spaces with water. Some stamping
is to be recommended. Then another layer of 3 to 4 in. of rock is
added, the spaces filled with sand with the help of water, and
stamped again. Then gravel, and as final layer sand, is used.
The top of the sand layer should come within an inch of the
width of a brick if the upper edge of the brick is to be flush with
the bottom of the working-door frame. This done, an inch
layer of clay is spread with the trowel. A straight-edge and
level ought to be used. Then time has to be given for the clay
to dry, while work is being done on other parts of the furnace.
If the furnace is very long several cross walls should be made,
well connected with the side walls.
When the clay has sufficiently dried, the brick pavement is
laid. The hardest bricks are selected for it. They are set on
the narrow edge and with their long side parallel with the fire-
bridge, which makes the hoe slide easily over them. They can
either be set dry and the spaces filled with sifted sand, or they
can be set in clay, but always as close together as possible. The
hearth next to the fire should always be set dry in sand to permit
the bottom to expand without bulging. For the binding rods
which pass under the hearth, channels should be made of bricks,
so that these rods will be kept cool by air and can be easily
inserted or withdrawn.
As regards the roof r, the arch over the hearth should be made
pretty flat in order to spread the flame, but care should be taken
not to go to the extreme. A rise of the arch of 14 to 17 in. over a
10 ft. wide hearth is about proper; to give such an arch only a rise
of 5 to 6 in. is a great mistake if any durability of the furnace is
expected. It has to be considered that in the heat the hearth
will expand, and if the joints between the bricks are not very
close and carefully made, the arch will soon cease to be an arch,
but become a flat plane which in course of time even becomes an
inverted arch, and soon will cave. The arch is usually made
9 in. thick. If square bricks are used, which is generally the
case in remoter mining districts, it is much better to throw two
4J-in. arches one on top of the other than to make only one arch
by setting the bricks on their narrow edge, because then the
joints are so much wider on the outside than on the inside that
they have to be filled with clay and brick chips. This filling offers
54 HYDROMETALLURGY OF SILVER
very little resistance and makes the arch weak. The foot of the
arch should rest on the long sides of the furnace and extend also
over the fireplace. It is not necessary to cover the latter with a
separate cross arch.
With regard to the hight of the roof above the hearth bottom
it can be considered, as a rule, that the hight should be the greatest
on the two hearths next to the fire, and then gradually diminish
to the flue end. In ascertaining this distance one has to be
guided by the character of the ore. For certain ores the roof of
the hearth next to the fire has to be made much higher than on
the following hearth, as is shown in Fig. 4, where the roof of
this part of the furnace is made 5 ft. above the hearth bottom,
or 2 ft. higher than that of the adjoining hearth. This is accom-
plished by a step of 2 ft., which, however, is an exceptional case.
Usually a long furnace will do good work if the highest part of
the arch of the finishing hearth is made 30 to 36 in. above the
bottom and then slopes down to 24 in. at the flue end. This
will allow live air to enter between the ore and the fire gases.
The spring of the arch, as stated, should not be made less than
14 to 17 in. These dimensions are given for highly sulphureted
ore, and have to be varied according to the character of the
ore.
In the endeavor to reduce the consumption of fuel to a mini-
mum we often find that other more important points have been
sacrificed to this one object. One of them is to make the arch
and sides too low. The flame, being pressed down by the low
arch, comes in contact with the ore, and exercises a reducing
action, which is adverse to the principle of chloridizing roasting.
The space between the ore and the roof is almost completely
filled with gases from the combustion of the fuel, and the live
air has no opportunity to enter deeper into the furnace. The
little that enters through the working doors is forced to the side,
doing some good work on its way to the flue, but the main portion
of the ore depends for its oxidation on the small volume of unde-
composed air mingled with the fire gases. Thus a furnace may
have plenty of draft but not enough air. This defect in the
construction is felt still more if in a long furnace an ore is to be
roasted which cannot stand a high heat without caking or suf-
fering a great loss of silver by volatilization, because it is not
possible to keep the heat on the finishing hearth as low as required
REVERBERATORY FURNACES WORKED BY HAND 55
and at the same time insure the working on the more remote
hearths to the best advantage.
Instead of using a wooden center to build the arch, quite
frequently damp sand is used for this purpose. The sides are
finished first, including the skew back, then the center line is
marked on the bottom of the furnace and wooden sticks set up
on that line at distances of 3 to 4 ft. and kept in position by
some moist sand. The length of these sticks must correspond
with the hight of the roof at their respective places. All the
working doors and other openings are closed with boards and
then the furnace is filled with moist sand, or tailings. This done
the shape is given to the arch with a straight-edge and trowel.
The finishing is done with a thin layer of lime mortar, on top of
which the arch is built. Clay should be used as mortar for the
arch. After several days the sand is removed, but not until
buck stays and binding rods are placed in position; otherwise,
when by removing the sand the weight and side pressure of the
flat arch are thrown on the sides, they may give way and cause
the arch to cave.
The charge hopper h, is made of stout sheet iron and should
be large enough to hold a full furnace charge, which consists of
1500 Ib. to one ton of ore. Not to burden the arch with this
weight, the hopper is flanged with angle iron around the rim
and hangs on a framework which rests on two sides of the furnace,
as shown in Fig. 4. These hoppers are usually made square, and
the narrow end, which is provided with the slide, S, corresponds
with a cast-iron extension which passes through the arch. This
piece of casting should be shaped to correspond with the circle of
the arch, so it will act as a keystone. The hopper is filled by
means of dumping cars running on an elevated track.
The flue-hole should always be made in the end wall of the fur-
nace, and not in the roof, because the flue (see e, Fig. 4) on top
of the furnace is much in the way. It is of the greatest importance
to have ample draft in the furnace. The draft in the furnace
depends on the suction power of the stack and on the size of the
flue-hole. If the latter is made too small the gases will be
throttled and the furnace will smoke. No good roasting can be
performed in such a furnace; and as the draft is such an important
factor in roasting, the flue should be made full large and be
provided with a damper to regulate the draft. Dampers are
56 HYDROMETALLURGY OF SILVER
indispensable, especially if a number of furnaces are worked by
the same chimney. The furnaces next to the flue leading to the
chimney will receive an excess, while those situated farther away
may not receive sufficient draft.
I found that through insufficient supply of air not only the
chlorination suffers, but the loss of silver also increases. If the
roof of a furnace is of proper hight to permit air to enter between
the ore and the gases, and the fire-bridge and the back of the
furnace are provided with air-ducts, the interior of the furnace can
be observed, which makes it easier to regulate the draft. The
draft must be so regulated that the fumes evolved are kept in
motion in all parts of the furnace. If they stagnate around the
ore, or if the furnace assumes a nearly uniform heat throughout
the entire length, it is always a sign of insufficient draft, and if
the draft is not increased the result will invariably be a high loss
of silver and a low chlorination. On the other hand, if the
flame coming from the fireplace becomes short and pointed and
travels very swiftly, it is a sign of too much draft, which causes
an unnecessary consumption of fuel and may cool the furnace too
much for it to do good work.
If the flue opening in the end wall of the furnace is made
4 to 5 ft. wide, the sides 9 in., and the spring of the arch 12 in.,
it will answer for all kinds of ores.
For the working door /, (Fig. 5) the best design for the
reverberatory furnaces is represented by Figs. 7 and 8. It is
18 in. wide and 10 in. high. Its design is ingenious and simple.
Each side is in the form of an angle, which enables the laborer
to reach through one door with his hoe all points of a 10-ft. hearth,
while if the two sides were straight it would require two doors
to each hearth to reach all the points. The top and bottom plate
of the frames extending beyond the sides, and the angular shape
of the latter, make it possible to cement solidly the door frame
into the wall, without the use of any anchor-bolts, which are not
of much use anyway, as they invariably work loose in a short
time and then loosen the bricks around the door. A movable
iron plate serves as door.
Each door is provided in front with a 2J-in. roller, the object
of which is to facilitate the working of the charge. The long
handle of the tool rests on it and in stirring it revolves, and thus,
while taking the weight of the tool, lessens the friction. The
REVERBERATORY FURNACES WORKED BY HAND 57
full advantage, however, of this arrangement is only obtained
when the movement of the hoe is at right angles with the roller.
As soon as the angle is changed the hoe will partly slide, or even
will slide altogether, while the roller stops. A far better contri-
vance is that designed by G. Kiistel, and shown in Figs. 9 A, 9 B
and 9 C. Instead of the long roller he provided the door frame
FIG. 7. — PLAN OF WORKING DOOR.
FIG. 8. — ELEVATION OF WORKING DOOR.
with a grooved wheel, which rests in a fork movable on a pivot.
The groove is large enough to take the handle of the hoe. In
working, the hoe moves easily to and fro on the wheel, which,
on account of the pivot in the frame, will turn to any direction
in obedience to the hoe and revolve with the same ease.
When the furnace is built, it has to be dried carefully with a
slow fire for several days. It is well to keep a small fire on each
hearth, so that the whole furnace will get about the same heat.
If fire is kept on the fireplace only it will take a very long time
to dry such a long furnace.
In starting the furnace it is best to charge each hearth through
the working door with crude ore. Each charge will have to
remain on the same hearth until the charge on the finishing hearth
58
HYDROMETALLURGY OF SILVER
is chloridized, which will be quite a number of hours if the ore is
highly sulphureted, but after the first charge is drawn and the
others are moved forward one hearth each, the time required
on the finishing hearth will be less, and soon the whole furnace will
be in good working condition. The discharging is done through
the working door of the finishing hearth, into wheelbarrows.
Shortly before discharging commences the charge of each suc-
ceeding hearth is raked into a pile on the forward half. As soon
as the finishing hearth is discharged the forward movement of
the charges begins, and when done a new charge is dropped on
the first hearth. On those hearths on which the oxidizing takes
FIG. 9 A. —THE KUSTEL WORKING
DOOR.
place the charges should be worked diligently to expose the ore
as much as possible to the oxidizing action of the air. This not
only shortens the roasting time but more sulphuric acid is formed,
producing a better sulphating of the ore. When the ore becomes
woolly, that is, when the chloridizing period sets in, much less
stirring is required. On the finishing hearth the ore should be
raked to a thick layer in the center of the hearth, the space next
to the fire-bridge and the back wall being kept clear. From near
the working door the ore should also be pushed farther into the
furnace, to prevent the cooling of that portion of the charge.
REVERBERATORY FURNACES WORKED BY HAND 59
A thick layer diminishes the loss of silver by volatilization. On
this hearth the ore should be stirred only a few times. If the
salt is to be added in the furnace, it is soon found by observation
and assays on which hearth this has to be done to obtain the best
results. Always consider both chlorination and volatilization.
With regard to the fire the operator has to be entirely guided
by the temperature required on the finishing hearth. If the ore
is such that it loses much silver if exposed to too high a heat,
which can be ascertained before the furnace is built, then, in
order not to reduce the working capacity of the furnace by keep-
FIGS. 9B and 9 C. —DEVICE FOR WORKING DOOR.
ing so low a heat as the ore requires, the roof of the finishing hearth
has to be made much higher above the hearth. This is best
done by dropping that hearth a step lower than the other hearths.
The level of the fireplace, however, should not be dropped too,
but should be put in proper position for the other hearths,
which, of course, will make a higher fire-bridge for the finishing
hearth.
Three men are sufficient to attend a furnace 50 ft. long, with
one helper for each two furnaces. When a charge is drawn, two
of the men and the helper wheel the roasted ore to the cooling
floor, while the third man pulls the charge. The fire is cared for
by the man who attends to the two hearths next to the fire
60
HYDROMETALLURGY OF SILVER
on which the ore does not need to be worked so frequently as on
the other hearths. The removal of the ashes is done by the
helper. The wood is supplied by one yard man to all the fur-
naces; likewise the hopper is filled with ore by one man for
all the furnaces. The roasting capacity of a 50-ft. furnace
is about 8 to 9 tons in twenty-four hours.
THE TWO-STORY LONG FURNACE
As a further improvement in the hand-worked reverberatory
furnace we have to consider the two-story furnace. Instead of
building the furnace in one direction, say 40 ft. long, it is built
FIG. 10. — LONG REVERBERATORY FURNACE, TWO-STORY.
FIG. 11. —LONG REVERBERATORY FURNACE, TWO-STORY.
in two stories of 20 ft. hearth each. There is by this method not
so much loss of heat by radiation. The hot arch of the lower
story warms the bottom of the upper hearths, and a new charge
is more quickly ignited than if the hearths are all on one level.
Figs. 10, 11 and 12 illustrate the construction of such a furnace.
REVERBERATORY FURNACES WORKED BY HAND 61
The lower two hearths are 11 ft. long each, while the upper two
are 10 ft. each. A number of these furnaces were erected by me
for the Mexican Santa Barbara Mining Company, to roast the
heavy zinc-lead ores of the San Francisco del Oro mine. In this
furnace 10 tons of ore were roasted in twenty-four hours. The
oxidizing was done on the upper two hearths. Through an open-
ing in the bottom of the second upper hearth the charge was
18 'Flue
FIG. 12. — SECTION THROUGH A B
(Fie. 10).
dumped into the first hearth of the lower furnace. The salt was
added in the upper hearth just before dropping the charge into
the lower hearth. In this way the salt was well mixed with the
ore. The construction of these two-story furnaces permits the
insertion of an auxiliary fireplace across the width of the hearth,
which can be used to advantage, because the heat is spread over
the whole width of the hearth, which is not the case in a long
furnace, where, if an auxiliary is used, the fire ha& to enter the side
of the furnace. In roasting the San Francisco del Oro ore the
auxiliary fire was used only for a short time after a new ore
charge entered the furnace, for the purpose of reducing the cooling
effect of the cold charge on the other charge, which was in the
state of roasting, and to ignite the fresh charge more quickly.
The fire was stopped as soon as the fresh charge showed the
sulphur flame. The gases from the lower furnace entered the
upper furnace through two flues, one on each side of the fire-
bridge proper, so the rising gases did not interfere with the flame
of the auxiliary fire.
At Sombrerete, Mexico, when it became necessary to rebuild
two of the 40-ft. furnaces, I changed them into 20-ft., two-story
furnaces, and increased thereby the working capacity of each by
nearly two tons per day.
VIII
MECHANICAL ROASTING FURNACES
IN this chapter will be included only such mechanical roasting
furnaces as are specially adapted for chloridizing roasting. There
are two classes of such furnaces: one in which the ore is roasted
in charges; and the second in which the roasting is continuous,
that is, furnaces in which at one end a continuous stream of
raw ore enters, while at the other a continuous stream of roasted
ore leaves the furnace.
As stated above, in no mechanical furnace can the process
of roasting in all its stages be so well controlled as in a rever-
beratory furnace worked by hand, and, therefore, their applica-
bility is much more limited to certain classes of ore. However,
if applied to suitable ores, they do very good work, and, where
labor is expensive, are more economical. In Mexico, where labor
is cheap, the mechanical furnaces proved successful only in
exceptional cases. All mechanical furnaces are connected with
more or less machinery and require frequent replacement of the
wearing parts. These parts of machinery and castings are sent
from the United States, and are rather costly by the time they
land in some remote mining place in the mountains. This,
however, would not be so important a factor if it were not for
other inconveniences connected with it, as the long time it takes
to get these parts, and, perhaps the sudden breakdown of parts
of which duplicates may not be on hand, requiring quite a long
shutdown of one of the furnaces, which always represents a large
percentage of the roasting capacity and with it of the producing
capacity of the works.
The erection of mechanical furnaces for ores for which they
were not suitable has caused many serious failures.
62
MECHANICAL ROASTING FURNACES
63
(1) MECHANICAL FURNACES FED BY CHARGES
(a) The Bruckner Revolving Furnace. — This ingenious device
of a chloridizing furnace was successfully introduced by its
inventor, Mr. Bruckner, in Colorado, in 1867. The furnace con-
sists, as illustrated by Figs. 13 and 14, of a cylinder of boiler iron,
the ends of which are closed save a circular central opening on
each end. Two circular tracks are fastened around the cylinder
placed at even distances from the ends. With these tracks the
cylinder rests on four strong wheels. A revolving motion is
imparted to the cylinder either by friction, in which case one of
FIGS. 13 and 14. — BRUCKNER ROASTER.
the four wheels is made to revolve, or the motion is imparted by
pinion and cogs, in which case a cast-iron ring with cogs is fas-
tened to the cylinder. At one end of the cylinder is placed a
fire-box, the throat of which corresponds with the central end
opening of the cylinder, while the opposite opening corresponds
with a circular hole in the flue. Four doors are placed diamet-
rically opposite, two on each side. These doors serve for charging
and discharging the furnace. Above the furnace is placed a
hopper large enough to hold a charge of ore. The hopper has
two outlet spouts, each provided with a slide, which correspond
with the furnace doors. The shell as well as the ends are lined
64 HYDROMETALLURGY OF SILVER
with bricks. Provision is made in the driving mechanism to
regulate the speed from one revolution in one minute to one
revolution in three minutes. Some of the furnaces are so con-
structed that the two ends of the cylinder are slightly contracted
in order to facilitate the discharging of the ore, but this compli-
cates the lining and reduces the capacity, and is actually not
necessary, as a straight cylinder discharges very nicely, and if
a little of the charge does remain in the furnace, it helps to heat
the new charge.
The manipulations of this furnace are as follows : The two
doors of one side are opened and the furnace revolved until the
doors come right under the two spouts of the hopper, when the
two slides are withdrawn and the charge allowed to drop into
the furnace. Then a quarter turn is given to bring the doors in
proper position for the man to close them. They are fastened as
tight as possible by means of a wedge. Though the lid and the
flange of the cast-iron door frame are faced it is not possible,
especially if the furnace has been in use for some time, to close the
door perfectly tight, and when the furnace revolves, some ore
will leak out through each door. This takes place only as long
as the charge is crude and stops when actual roasting is in progress.
This leakage is the more annoying as the ore is crude and mixes
with roasted ore underneath the furnace. To prevent this leak-
age the joint of lid and flange should be plastered from the out-
side with clay, or better with paste of sifted wood ashes and
salt. Before charging the furnace should be well heated, and
during charging the draft checked so that not too much dust is
carried into the flue.
In the beginning a strong fire is kept, but as soon as the sul-
phides are well ignited the fire is allowed to go out. The heat de-
veloped by the oxidation is considerable, and if further increased
by the fire a caking and balling of the ore would take place.
These furnaces are usually 6 ft. in diameter inside the lining and
16 ft. long, and take a charge of 5 to 5J tons of ore. Usually two
charges can be roasted in twenty-four hours. About three to
four hours of each charge the furnace can run without fire, then
the salt is added and the roasting continued with a moderate
fire. A hole back of the flue permits the observation of the
temperature and the taking of samples with a long-handled scoop.
During chloridizing the charge in the furnace assumes an inclined
MECHANICAL ROASTING FURNACES 65
position up to 45 deg., the weight of which acts against the direc-
tion of the motion, and if the clutch is thrown out in order to
stop the furnace, this weight will pull the furnace back nearly a
quarter of a turn. To open the doors to add the salt or to dis-
charge the furnace, it is necessary that the furnace should be
stopped at a position convenient to the roaster man. In order
to accomplish this the end of an iron bar is pressed between the
track and the wheel at the side at which the movement of both
is outward. When the doors are in the right position the clutch
is thrown out, but the furnace is prevented from revolving back
by the iron bar, which becomes clamped in very tightly. By means
of a short-handle shovel the salt is introduced through the two
doors and well scattered over the whole surface of the ore. The
salt decrepitates violently and in this way becomes very evenly
divided. After the salt is added and the furnace revolves again,
the ore becomes woolly, as in the reverberatory, and assumes a
still more upright position.
If the ore is sufficiently sulphureted and the salt is added it
will not be necessary to start the fire again. There is enough
heat stored in the charge to finish the process of chlorination
(see remarks on chloridizing self-roasting), in which case the con-
sumption of fuel is very small. This mode of roasting is only
admissible if the ore is to be roasted for lixiviation. For amalga-
mation a second fire is indispensable, because the base-metal
chlorides have to be decomposed or volatilized.
When the charge is finished two cars are pushed under the
furnace, one for each door. They are large enough to receive the
full charge of the furnace. All four doors are opened, the lids
kept in position by a proper contrivance, and the furnace is re-
volved again. While the furnace is prepared for discharging, a
good strong fire should be started again, to heat it for the next
charge. The receiving cars are made narrow and long, so that
no ore is dropped beyond the rim of the car, because, especially
in the beginning, the ore will pass through the door over a
large arc.
The cooling floor is situated 5 or 6 ft. lower than the track
under the furnace, which track extends some distance on iron
trestle over the cooling floor. The body of the cars is shaped
like a hopper with bottom discharge, and closed by a slide which
is worked by a lever. As each car holds about 2J tons and is
66 HYDROMETALLURGY OF SILVER
hot, they are pulled out from under the furnace by means of a
windlass and chain.
The revolving motion of the Bruckner furnace should be
slow. There is not the least advantage in whirling the ore around
and around in the furnace. When new ore particles are brought
to the surface it takes some time to undergo oxidation; why not,
then, give them the required time to be in contact with the air be-
fore again immersing them under the surface? I found in some
works large Bruckner furnaces set to make two and even two and
a half revolutions per minute, whereas a speed of one revolution in
three minutes is ample. A Bruckner furnace, when charged,
weighs about 16 tons; this divided on four wheels makes 4 tons
to the wheel, a rather heavy weight, especially if we consider
that the whole of this weight is pressing against the space of con-
tact between the ring-track of the furnace and the surface of the
wheel, which is very small. The effect of this high pressure is
shown by the way the wheels and the ring-tracks wear. When
the furnace is revolving it can be observed that continually thin
scales of iron up to the size of a finger nail are dropping from the
face of the wheels and tracks. Now if two revolutions are made
in one minute, instead of one revolution in three minutes, the
wear will be six times as great, to say nothing about the greater
power which is required for the faster motion. If the capacity
of the furnace should be increased by it six times it would be
different, but this is not the case; a charge roasted at a high fur-
nace speed takes just as long to be finished as a charge does when
rotated at a moderately slow speed. Besides much increasing
the wear, a high speed can cause serious trouble if the ore has a
tendency to ball, in which case it may happen that the whole
charge is transformed into balls from the size of a pea to that of
a cocoanut, without leaving any fine stuff at all.
One of the main advantages of the Bruckner furnace is the
fact that the ore in this furnace is kept in a thick layer and still
permits a thorough oxidation and chlorination. As explained
above, the loss of silver by volatilization is much less if the ore
during roasting can be kept in a thick layer, because a large
portion of the evolving volatile chlorides is kept back in the ore
as by a filter, and consequently much less silver will be carried
away.
Under equal conditions, it will be found that an ore loses less
MECHANICAL ROASTING FURNACES 67
silver if chloridized in a Bruckner than in any other furnace in
which the ore is roasted in a thin layer.
Ores not suitable for the Bruckner furnace are those which,
on account of a large percentage of lead, antimony, etc., cake
easily, because by the continuous rolling of the ore any lumps
which may have formed cannot be mashed shortly after they
are formed, but increase much in size by the continual rolling,
which makes them hard and dense.
(6) The 0. Hofmann Improved Bruckner Furnace. — In the
Bruckner furnace the ore is exposed to a rather uneven heat.
The part next to the fire receives always the highest heat, while
that at the other end, 16 ft. away, will receive much less, in
some cases not even enough without overheating the fire end.
This is a defect of the furnace which is of no consequence if a
highly sulphureted ore is treated, which can be subjected to
self-roasting, because it creates ample heat by itself to become
uniformly hot through the entire length of the furnace; but it is
a defect much felt if ore poor in sulphur is to be roasted, which
has to receive nearly all the required heat from the fire, or an ore
which has to be roasted at a low heat, because it is apt to lose
much silver by volatilization, or lumps easily at a higher tem-
perature (see chloridizing roasting with steam, page 34). Mr.
Bruckner was aware of this defect, and he tried to remedy it by
inserting into the furnace a diaphragm made of cast-iron pipes.
This diaphragm was set at an angle of about 15 deg. to the axis.
It had a diagonal position extending through the whole length
of the furnace and was intended to move the ore from one end
to the other and back, in order to produce a uniform heating of
the charge. This device, however, did not give the satisfaction
expected and was soon abandoned, especially on account of the
inconvenience the diaphragm caused in cleaning the furnace from
the crust, which has to be done from time to time, and of the
short life of the pipes, though they projected through the shell
of the furnace to permit air to pass through.
Confronted with the necessity of obviating this defect of the
Bruckner furnace, because the rich ores of the Silver King mine,
Arizona, could not be roasted successfully in the Bruckner
owing to this defect, I changed the arrangement of the fur-
naces, inasmuch as I attached a fire-box and flue arrangement
to each end of the furnace, which enabled me to heat either
68
HYDROMETALLURGY OF SILVER
end of the furnace by changing at intervals the course of the
flame.
Figures 15, 16, and 17 represent the arrangement. Between
the fireplace proper and the furnace is situated the flue, extending
downward to the dust-chambers, Fig. 15. This flue is provided
with a damper. The other end of the furnace is provided with
exactly the same arrangement. The dust-chambers from both
ends are connected with the main flue leading to the stack.
Before they connect there is an additional damper, Fig. 16, on
each side, to make it sure that, if the dampers of one side are
closed, no draft passes through on that side.
=*n
iu-\-m
• ra1 1
j 1 ...-. • _
3H-
BP.J.. i ~n
FIGS. 15-17. — HOFMANN IMPROVED BRUCKNER FURNACE.
If the fire is kept at one side for some time the dampers of
that same side are opened and those of the other are closed,
and the fire started there. The flame now traverses the furnace
in the other direction, and the ore at that end will be exposed to
the same heat as the other was before. This changing can be
done at intervals to suit the character of the ore. The changing
of the fire does not cause any trouble, as the opposite fireplace
is still warm enough to ignite the wood when the change is made.
The results obtained with this furnace have been very satis-
MECHANICAL ROASTING FURNACES 69
factory. I obtained with it good results in roasting the ores of
the North Mexican Silver Mining Company, Mexico, which were
very poor in sulphur and could not be heated sufficiently at
the opposite end with a fire at one end only. With the double
arrangement quite satisfactory results were obtained.
If a number of these furnaces are built in a row the fires of
all of them have to be changed at the same time, which is neces-
sary on account of the dampers in the two wings of dust-chambers
and on account of the man attending the furnaces. If the fires
are all on one side he will and can attend better and to more
furnaces than if the fires are partly on one and partly on the
other side.
There is quite a serious omission in the construction of the
regular Bruckner furnace, and that is that there is no provision
made for the admittance of live air into the furnace. Air being
such an important factor in roasting it is absolutely necessary
to have some means to admit air if the roasting is to be conducted
intelligently.
To keep the fire-door open for this purpose is not to be recom-
mended, as this shortens the flame. In the Hofmann furnace
provision is made for an air inlet by making the circular cast-iron
throat of the fire-box longer than usual and by leaving in the
lower half of the same a sufficiently large opening for the air.
The size of the opening can be regulated by a hinged door and a
lever.
On the periphery of the furnace and near each end is a small
door for taking samples. These small doors are easy to handle,
and they can be opened and closed and a sample taken without
stopping the furnace, which is quite convenient.
The spouts of the hopper have to be made to stand pretty
high above the furnace in order to permit the passage underneath
of the door and the eye to which the door is keyed when open
for discharging. This causes considerable spilling of ore during
charging. To prevent this, the spouts are each provided with a
sliding sleeve kept in position by a lever and weight, Fig. 17.
When the furnace is stopped with the doors open for charging,
the weight of the lever is removed and the sleeve lowered until
it projects into the door. Then the hopper slide is pulled. This
arrangement permits very clean work.
The lining of the furnace has to be done very carefully. The
70 HYDROMETALLURGY OF SILVER
door frames projecting inside should conform with the circle of
the lining. Specially made bricks to fit the circle of the arch
should be used only; if not, the lining will soon come out. To
facilitate the work of lining it is well to rivet to the steel shell of
the furnace six angle-iron ribs the whole length of the cylinder,
which will divide it into six equal sections. The projecting part
of the rib should be 3^ in., so that when the bricks are laid the
latter project half an inch above the rib. The groove which is
formed by it is filled with clay. Thus each section forms an arch
for itself, the angles serving as skew-backs. Each should be
well keyed. In lining, the bricklayer can then bring the cylin-
der always in the most convenient position for his work, as there
is no danger of caving even if that part of the lining which was
made first comes to stand right above him. If, in course of time,
from some reason or other, part of the lining should become
defective, the bricks of the bad section can be taken out and
replaced without disturbing the other parts of the lining. The
ends should be lined first, so that the lining of the cylinder will
abut against it.
As the furnaces of this type when charged are very heavy,
the tracks with which they rest on the wheels should be made
very solid, allowing considerable iron for wear. At the places
at which the two tracks are to come a wrought-iron band an
inch thick and as wide as the track, including its two-flange pro-
jection, is to be strongly riveted to the shell. To this band the
track is fastened by means of tap-screws. Bolts will not answer
because, if the track is worn and has to be replaced by a new one,
it will be found that many of the nuts are so tightly roasted in
that the bolts will be turned off in trying to unscrew the nuts,
and to renew these bolts means the taking out of a part of the
lining, which ought not to be. If a tap-bolt breaks, the piece in
the wrought-iron band can be bored out and a new tap-bolt put
in without difficulty. It is best to have the track cast in about
four segments. The joints of the segments should not be square
with the track, but slanting, because the furnace will revolve
more smoothly and the chipping of the edges of the joints will be
much less.
As in any other roasting furnace, in course of time a crust is
formed, which has to be removed from time to time. This is
usually done by cooling down the furnace and having men remove
MECHANICAL ROASTING FURNACES 71
the crust by picks, shovels and other tools. When cold the
crust is quite hard, and the removal of it endangers and weakens
the lining. While hot this crust, however, is rather soft, which
suggested to me a method by which the furnace can be quickly
cleaned without cooling or shutting down the furnace. When
the furnace needs cleaning, after it has been discharged, a
charge is made up of bricks, fire-bricks if possible, and the furnace
revolved while a very strong fire is kept. The heat softens the
crust, and the bricks in moving in the furnace gradually shave
off the crust without injuring the lining. This is done in a few
hours, and when the bricks and crust are discharged the furnace
is not only in shape but also heated to receive a new charge.
Besides this, the men are relieved from an unhealthful job.
There are other roasting furnaces which properly belong to
the class of mechanical furnaces fed by charge, like the revolving-
hearth furnaces; but they never were used much in actual prac-
tice, they are not very convenient, and are of small capacity, so
that I can without hesitation leave them undescribed.
(2) MECHANICAL ROASTING FURNACES WITH
CONTINUOUS FEEDING
To this class of furnaces belong the O'Harra, the Ropp the
Howell-White and the Stetefeldt. The Brown and Pearce fur-
naces, while they are excellent for oxidizing roasting, are not
quite suitable for chloridizing roasting.
(a) The O'Harra Furnace. — This furnace was the first mechani-
cal furnace with continuous feeding devised on the Pacific slope.
It is very ingeniously arranged and is, in improved form, still in
use. In the following I give a description of this furnace by the
late G. Kiistel, who had much experience with it:
"This furnace was first tried in 1862 or 1863 in Dayton,
Nevada, and later three of them were built in Flint, Idaho. The
main feature of this furnace is the endless chain to which two
oval rings are attached, the wings being as wide as the cross-
section of the hearth. To these rings are fastened the plows or
shoes by which the ore is gradually pushed forward. The hearth
of the furnaces built in Flint were 104 ft. long and nearly 5 ft.
wide. Eighty feet of this hearth were covered by an arch 12 in.
high; attached to it were three fireplaces — two on one side, and
72 HYDROMETALLURGY OF SILVER
one between the two on the other side. At one end was the feed-
ing hearth, which was not covered by the arch, and on which the
ore was continually delivered from the stamp battery by mechani-
cal contrivances. The motion of the ore was effected by an endless
chain, passing over two chain wheels, one at each end. To this
chain two oblong flat rings were attached, each provided with
eight shovels or plows, so arranged that while one of the rings
shoveled the ore toward the center line, the other pushed it back
again toward the sides every three or four minutes (or in shorter
intervals if more ore is charged). The ore not only changed its
place to the right and left, but it also moved forward by degrees,
so that in course of six hours from the beginning it commenced
to be discharged, passing 18 ft. over a cooling hearth. On both
ends of the furnace were iron doors hung on hinges which were
opened by the ring every time it passed.
"The whole plant at Flint was arranged to work automatically.
The five batteries, of five stamps each, had on both long sides end-
less screws, by which the crushed ore was forwarded in proportion
as it discharged from the battery, and dropped into an eleva-
tor. Having been lifted about 15 ft., it was conveyed again by
endless screws along the feeding hearths of all three furnaces.
The discharge of this conveyor was so regulated that each feeding
hearth received an even part of the ore. The ore mixed with
5 per cent, of salt was spread on iron plates behind the batteries
(heated by the hot gases from the furnaces, conveyed through
the flue and under the plates). When charged into the battery,
the ore was not further handled till it came out of the furnace
perfectly roasted.
"There is only one obstacle connected with this mechanical
furnace. The shoes or shovels, touching the sides of the furnace,
wear off by degrees, leaving a space which is taken up by the ore.
This part of the ore along the wall hardens and increases in
amount in the furnace till new shoes are put in. By these the
crust of one-half to three-quarters of an inch thick is broken off
and carried out. From the Rising Star ore these crusts contain
nearly as much silver chloride as the well-roasted ore; they are,
nevertheless, disagreeable, but some means might be devised by
which this inconvenience could be avoided.
"The ore from the Rising Star mine at Flint contained
argentiferous fahlerz, miargyrite, ruby silver, zinc blende, galena,
MECHANICAL ROASTING FURNACES
73
iron pyrites, and sulphide of antimony. At an average the ore
contained between 69 to 77 ounces silver per ton and some gold.
The gangue was quartz. It was crushed through sieves with
forty holes to the inch, together with 5 per cent, of salt. The
ore in the furnace, when reaching the first fireplace, commenced
to roast oxidizingly. Between this fireplace and the second, which
was on the other side, the chlorination began at an increased
heat. Between the second and third fireplace the chlorination
was finished at a high red heat. Although not more than 5 per
cent, of salt was added, the roasted ore contained about 90 per cent,
of the silver converted into a chloride. The gases, containing
free chlorine and chloride combinations emitting chlorine, coming
in contact with the surface of the ore while passing over it for a
FIG. 18. — O'HARRA FURNACE.
space of eighty feet, have a chloridizing influence on it, replacing
thus a certain amount of salt.
"The capacity of the three furnaces was more than twenty
tons. Each one could easily treat ten tons of the Rising Star
ore in twenty-four hours. The roasted ore was treated by
amalgamation in pans."
Mr. Kiistel continues:
"O'Harra's furnace is now greatly improved (Fig. 18). It is
built in two stories, so that when the chain comes out of the
lower hearth it turns into that of the upper story. The chain is
heavy and there is no trouble whatever. Although the chain
in its course through the red-hot furnace is exposed to red heat,
74 HYDROMETALLURGY OF SILVER
it nevertheless does not become so hot as to suffer any injury
from it; there is a wooden framework on either end of the furnace
over which the chain and plows move in the open air, which
prevents them from getting too hot. The furnace, of the latest
construction, is eight feet wide and from forty to a hundred feet
long, with four fires, two on each side, directly opposite each
other. The first two fires, where the ore comes in contact there-
with, are divided, so that one-half of the flame goes direct to the
lower, and the other half to the upper hearth, through an opening
in the arch of the fire-chamber. There is a fire-clay damper to
regulate the flame. The other two fires are opposite each other,
so that the heat is uniform over the whole hearth. These last
two fires are regulated by the ash-pit dampers.
"The endless chain has two triangular frames, with plow-
shoes on each side. The cooling space is built in proportion to
the length of the hearth. A furnace of this kind, fifty feet long
and eight feet wide, can roast from thirty to forty tons in twenty-
four hours, with the help of only two men, consuming about two
cords and a half of wood. The chlorination runs up to 90 and
95 per cent.
"The working of this furnace is not expensive, as two men,
one at daytime the other at night, can attend the roasting of
forty tons.
"A remarkable feature of O'Harra's furnace is the very small
amount of dust that is carried off by the draft. Another pecu-
liarity of the furnace is for drying ore in pieces the size of a man's
fist. One furnace near Shasta, California, dries 40 tons of ore in
twenty-four hours, at a small expense. "
The construction of this furnace does not offer many facilities
for regulating the process, and therefore it can be used for chlo-
ridizing roasting only for a particular class of ore. To roast in
this furnace ores which are apt to lose much silver by volatilizing
would be rather risky. Moreover this furnace is not suitable for
roasting ores which require the addition of salt after the sulphating
period, at least not in its present construction. However, this
furnace is well adapted for quite a variety of ores, especially those
which do not cake easily and contain a considerable quantity of
iron pyrites.
(6) The Ropp Furnace. — It is apparent that by dragging the
plows on the bottom of the furnace, as in the O'Harra furnace,
MECHANICAL ROASTING FURNACES
75
the bottom and the rabbles or plows will suffer much wear, and not
less the chain, especially when highly sulphureted ore is roasted.
Alfred von der Ropp has obviated these obstacles by a very
ingenious construction of his furnace. Figs. 19 and 21 B repre-
sent a horizontal section, and Fig. 21 A an elevation of his fur-
nace, while Fig. 20 is a cross-section through fireplace and hearth.
FIG. 19.— HORIZONTAL SECTION OF ROPP FURNACE.
It is a one-story straight hearth furnace, 105 ft. Jong and 11 ft.
wide in the clear. The hearth is longitudinally divided into two
even parts by a slot, which extends the whole length. The two
sides of the slot project above the hearth surface to prevent the
ore from falling into it. The slot communicates with a tunnel, E,
underneath the furnace. This tunnel is provided with a track
FIG. 20. — CROSS-SECTION OF ROPP FURNACE.
on which travels a four-wheeled truck or carriage. To this truck is
fastened an iron arm which extends through the slot into the fur-
nace, and to the end of which is attached a cross arm extending
over the whole width of the hearth. This cross arm is provided
with adjustable rabbles, which are designed not only to stir the
ore, but also to move it gradually forward. An endless wire
76
HYDROMETALLURGY OF SILVER
MECHANICAL ROASTING FURNACES 77
rope moves the truck and with it the rabble arrangement. The
truck, after passing underneath the furnace, turns a curve and
returns on the outside to the other end of the furnace, where it
enters the furnace again. On this track run four carriages or
trucks, which are placed at even distances from each other.
They are fastened to the wire rope. The rabbles are so arranged
that if the preceding rake turns the ore toward the right the
following one will turn it to the left. The rabbles can be set
more or less slanting as circumstances require, and also can be
lowered or raised. The tunnel underneath the furnace is high
enough for a man to enter and to pass underneath the truck in
case any repairs should be required. Three fireplaces provide
the necessary heat. They are all on one side of the furnace and
are constructed as shown in Fig. 20. The feeding of the ore is
done automatically by Challenge feeders, at one end of the fur-
nace. At the opposite end the roasted ore is brought out, each
rake pushing forward a certain amount and dumping it into two
iron cars. The two ends of the furnace are closed, like the
O'Harra furnace, by swinging doors, which are opened by
fenders attached to the rake and which drop back in position
after the rake has passed. Numerous doors on both sides of
the furnace serve for regulating the admission of air and give
access for cleaning the hearth.
The driving mechanism is of simple construction. Two bevel
gears transmit the power to a horizontal sheave around which
the steel wire rope travels. A similar sheave is arranged on the
opposite end of the furnace. The rope is entirely outside the
furnace and is therefore perfectly protected from the heat, and
so are the carriages. The rakes returning on the outside of the
furnace for so long a distance have sufficiently cooled at the time
when they enter the furnace again.
It is claimed that the furnace has a capacity of 36 tons in
twenty-four hours roasting an ore which contains 20 per cent,
sulphur, 8 per cent, lead and 17| per cent. zinc. This furnace is
also excellently adapted for oxidizing roasting of iron pyrites
and copper matte.
With respect to the roasting conditions of this furnace they
are the same as prevail in the O'Harra, but the mechanical con-
struction is much superior to that furnace.
(c) The Howell-White Furnace, — About the simplest and
78
HYDROMETALLURGY OF SILVER
cheapest continuous mechanical furnace, and at the same time
perhaps the most effective, is the • Howell-White. It is very
simple in construction, requires little repair and very little power
to run it, while its capacity is quite large.
This furnace is a revolving cylinder, open on both ends. One
of them is connected with the fireplace while the other enters
into the flue. The cylinder is made of cast iron in flanged sec-
tions, which in setting up are bolted together. The two sections
nearest to the fire are lined with bricks to protect the iron, and
in order to have the inside diameter, after lining, the same through
the whole length of the furnace, the diameter of these two sec-
tions is 10 inches larger than the balance. The furnace rests on
the top of five wheels all in one line and properly divided, as
can be seen in Fig. 22. It is kept in place by four rollers which
FIG. 22. — HOWELL-WHITE FURNACE.
touch the furnace half-way up its diameter, which are kept there
by strong, stiff iron frames. The furnace is made to revolve on
top of the wheels to diminish friction and consequently to make
it turn more easily. It has a slight inclination toward the fire.
The flue end of the cylinder is provided with a flange to prevent
the ore from falling into the dust-chamber. The feeding is done
through an inclined cast-iron spout, which passes through the
arch of the dust-chamber and extends into the furnace to within
five or six inches of the periphery. The feeding is done from a
hopper by means of a worm, the speed of which is adjustable.
To the inside of the unlined part of the cylinder are riveted
several iron ribs in the shape of a spiral. The object of these ribs
is to lift the ore, when the furnace is revolving, and to shower it
through the flame. As this causes much dust, which is carried
away by the draft, an auxiliary fireplace is arranged at that end,
MECHANICAL ROASTING FURNACES
79
the flame entering the dust-chamber right under the end of the
furnace.
The roasted ore leaves the furnace gradually at the fire end
and drops into a vault underground but conveniently accessible
from the cooling floor. For this purpose a space about 12 in.
wide is left between the fire-bridge and the furnace end. The
FIG. 23.— HOWELL FURNACE, DISCHARGE END AND ORE-VAULT.
revolving speed of the furnace is adjustable; in most cases 2J to 3
revolutions per minute is sufficient. The forward movement of
the ore in the cylinder is caused by the revolving motion, and
the speed, therefore, regulates the amount of ore which passes
through and the amount which is retained in the cylinder while
in operation. Fig. 23 represents a longitudinal section showing
80 HYDROMETALLURGY OF SILVER
the relative position of fireplace and furnace, the ore-drop and
the ore- vault.
The constant showering of the ore, brought up by the ribs,
through the draft causes a separation of dust and sand, the dust
being carried into the dust-chamber. The amount of dust de-
posited in the chambers as compared with that of the roasted ore
dropping into the vault is from 30 to 50 per cent. The dust is
well roasted, provided proper attention is paid to the auxiliary
fire and that it is kept strong enough. This, however, can only
be done if the fire is intrusted to an extra man. If it is made
part of the duty of the man attending the fire at the discharge
end, the auxiliary will always be more or less neglected. One
man at each end can attend to three furnaces.
There are ores, however, which will not stand entering the
sudden heat of the auxiliary without caking. Even if the dust
is roasted well by the auxiliary fire, the formation of so much of
it is very annoying and inconvenient. Actually there is no need
to make so much dust with this furnace.
This excessive amount of dust is caused by the ribs, which
produce a shower of ore through the swiftly moving gases. I
found that these ribs are not necessary. There is no accumulation
of ore in this furnace; in fact there is only a comparatively small
stream of ore passing through, even if roasting is conducted at
the rate of 30 tons in twenty-four hours. By the revolving
motion of the furnace, this thin layer of ore is made to expose
continually new particles to the action of heat and air, and the
ore at the discharge end will be found just as well roasted with-
out as with ribs; in fact better, because it will contain a larger
percentage of fine material. Besides, the furnace after being
incrusted offers a rough surface to the ore, which prevents its sli-
ding swiftly and spreads it over quite a large surface. By remov-
ing the spiral ribs from the Howell furnaces of the Cusihuiriachic
Mining Company, Mexico, I very much diminished the formation
of dust ; in fact so much so that the maintenance of the auxiliary
fire was not justified, and was abandoned altogether. The dust
from right behind the furnace was removed twice a day and ele-
vated to the feed-bin, thus mixing it with the crude ore and
feeding it into the furnace again.
To have an elevator between each two furnaces, which can
be made to discharge into either of the two feed-hoppers, is very
MECHANICAL ROASTING FURNACES 81
convenient, not only for elevating the dust from the chambers,
but also to elevate the ore-sweepings, which always accumulate
around a roasting furnace.
This furnace radiates considerable heat, which affects the
driving-belt and shortens its life; it is, therefore, much better
to drive the furnace by means of sprockets and link chain.
It ought to be mentioned that there are two vaults beneath
each furnace which can be filled alternately by turning a wing
which is placed right under the roasted ore drop. On each side
of this drop there should be a small door to permit the entrance
of air into the furnace and to give access for tools in case it is
necessary. »
If one of the vaults is filled with ore, it is advisable not to
discharge it until the other is nearly filled. This gives the ore
an opportunity to improve in chlorination.
When the crust in the furnace becomes too thick it can be
easily removed by inserting a number of bricks at the flue end.
In revolving the bricks will shave off the soft crust and bring
it out. When the bricks are charged the feed has to be stopped
for a while; but as soon as they have moved away 4 or 5 ft.
from the end, charging can be commenced again.
(d) 0. Hofmann's Modified Howell Furnace. — The Howell
furnace is a very efficient furnace; its first cost is small as com-
pared to its roasting capacity; it does not need many repairs, and
requires but very little manual labor. The furnace does not
carry much ore at a time, and is therefore not excessively heavy;
the way it is arranged, revolving on top of wheels, the friction is
much reduced, so that the furnace revolves very easily and not
more than two, perhaps two and a half, horse-power is required.
All these favorable qualities make the Howell furnace a very
desirable one; but it has the drawback, in common with all the
continuous furnaces, that the salt has to be mixed with the ore
before it enters the furnace. We have seen above that there
are ores which cannot be chloridized unless the salt is added
later during roasting, and as the construction of the Howell
furnace does not permit of this, it makes it unfit for such ores,
which is to be regretted considering the numerous advantagous
features of this furnace.
In my metallurgical investigations of the heavy argentiferous
zinc-lead ores of the San Francisco del Oro mine, Chihuahua,
82 HYDROMETALLURGY OF SILVER
Mexico, I was confronted, besides other roasting questions, with
the problem as to the best method of chloridizing these ores in a
Howell furnace. By mixing the salt with the ore in the stamp
battery the ore became sticky, incrusted the furnace rapidly, and
when it left the furnace consisted mostly of lumps without being
much chloridized. If the ore was charged without salt, it re-
mained loose and sandy, but it dusted so much that by the draft
an almost perfect separation of coarse and fine took place, the
latter being carried into the dust-chamber, while the coarse sand
dropped in the pit, but, on account of its large percentage of
lead and zinc blende, insufficiently oxidized, so that the salt,
which was added from time to time to the ore in the pit and
stirred, had but very little chloridizing effect. The best chlorina-
tion obtained was only 29 per cent. To diminish this separation
a small percentage of salt, from 1 to 2 per cent, was added in the
battery. The effect was remarkable; without balling or incrusting
the furnace the dusting was practically stopped. When the
balance of the salt was added in the pit a chlorination of 67 per
cent, was obtained.
Based on this observation a modification of the arrangement
in front of the furnace was made. Between the fireplace and
the discharge end of the furnace a shallow pit was inserted, which
was in communication with a reverberatory furnace 6 by 8 ft.
The bottoms of both were on the same level. The fireplace of
the reverberatory was only 24 in. wide. The gases from the
reverberatory passed through the pit and the furnace. When
a charge of about 1400 Ib. had accumulated in the pit the same
was pushed into the reverberatory, salt added and well mixed.
There the ore was kept until another charge had accumulated in
the pit. By this modification very satisfactory chlorination was
obtained. A very low fire was kept on both the fireplaces.
(e) The Stetefeldt Furnace. — This furnace consists of an
upright shaft 30 to 40 ft. high, which is connected near the top
with a descending flue. On top of the shaft is a feeding machine,
which showers the ore into the shaft. In the lower part are two
fireplaces opposite each other, so that the descending ore meets
the hot ascending gases. The principle on which the con-
struction of this furnace is based is at variance with that of any
other chloridizing furnace. The ore falls very finely divided
through a glowing atmosphere of chlorine, sulphurous acid,
MECHANICAL ROASTING FURNACES 83
oxygen and fire gases, metal chlorides and volatilized salt, and
the whole complicated reactions which take place in chloridizing
roasting are completed in the incredibly short time of a very few
seconds. Mr. Stetefeldt derived his idea from the Gerstenhofer
pyrites roaster, in which the ore is fed into a shaft, wherein
numerous shelves are so arranged that the ore drops from one
shelf to the other, resting on each for some time. The experi-
mental furnace Mr. Stetefeldt built was of such a construction
arranged so that fire gases were permitted to enter the shaft. The
shelves, however, caused much inconvenience by incrustation,
etc., and by observations he made during these experiments he
thought it justifiable to repeat the experiment without the use
of any shelves. The results were so gratifying that he, adhering
to this new principle, gradually developed his furnace to the
present much improved construction.
Fig. 24 represents a vertical section of this furnace. A, shaft;
G, returning flue; K, receiver or hopper forming the bottom of
the shaft; C, C, the two fireplaces; T, slit for the fire gases; U, air
ducts to produce a perfect combustion, and at the same time to
cool the fire arch; E, ash-pit, the iron door of which is provided
with a slide to regulate the air inlet ; 0, 0, doors for the insertion
of tools for cleaning the walls. The returning flue, G, is provided
with doors, R, which serve to clean it. D is an auxiliary fire-
place, which is constructed like the fireplaces of the shaft and
which serves to roast the large amount of flue-dust which this
furnace makes. Passing the chamber H the dust enters V and
the larger part of it settles in the bottom hoppers, /, /, from
which it is drawn into iron cars by moving the dampers S, S.
The rest of the dust is collected in a number of dust-chambers, Q,
which are connected with the chimney by means of a long flue.
P, P are doors for observation and cleaning. Below the shaft-
hopper, K, is the slide door L. By pulling this slide the accumu-
lated roasted ore drops into a large iron car. B is a cast-iron
frame with water-jacket on top of the shaft on which the ore-
feeding machine is placed. Above it is an ore bin (not shown
in the diagram) from which the ore is fed into the machine by
means of a worm.
The feeding machine is shown in Fig. 25. A is a cast-iron,
water-cooled frame placed on top of the shaft and provided with
the damper, B, which is withdrawn when the furnace is in opera-
84
HYDROMETALLURGY OF SILVER
tion, but which is inserted if the feeding machine stops for any
repair or for exchange of screens. C is a cast-iron grate to which
on the upper side is fastened the punched screen, D, which is
made of a steel plate with holes of an eighth of an inch. Above
the punched screen is placed a frame, E, to the bottom of which
is fastened a coarse wire screen, F, with about three meshes to
FIG. 24.— STETEFELDT FURNACE.
the inch, made of heavy wire. The frame, E, rests upon friction
rollers, G. The brackets, H, which hold the friction rollers, can
be raised or lowered by set-screws so that the wire screen can be
brought closer or less close to the punched screen. The brackets
K carry an eccentric shaft, by which an oscillating motion is
given to the frame E. To the brackets N are fastened transverse
stationary blades, 0, which extend close to the wire screen and
MECHANICAL ROASTING FURNACES
85
which can be adjusted by the nuts P. These blades keep the ore
uniformly spread over the screen when the machine is in motion.
The ore is usually crushed through a 40-mesh screen.
Before starting the feed the furnace has to be well heated,
which takes thirty-six to forty hours. After the speed of the
feeder is set and the proper temperature is ascertained by obser-
vation and numerous assays, the roasting itself requires but very
little attention beyond the maintenance of a uniform temperature.
FIG. 25. — FEEDING MACHINE, STETEFELDT FURNACE.
The capacity of the furnace is very large, roasting, according
to the size of it and the character of the ore, from 20 to 50 tons
in twenty-four hours. The consumption of fuel is rather small.
With one cord of wood about 8 tons of ore can be roasted. The
loss of silver by volatilization is less than if the same ore was
roasted at the same heat in another furnace, because, according
to Plattner, the loss of silver does not depend merely on the
temperature the ore is roasted at and the character of the ore,
but also on the length of time the ore is subjected to the roasting
temperature. In the Stetefeldt furnace the ore is roasted almost
instantaneously and does not suffer that part of the loss which
is caused by long heating.
This furnace is well adapted to roast ores not heavily charged
with sulphides, containing 5 to 8 per cent, sulphur, or even 10
per cent., if no, or only a small percentage of, galena and zinc
86 HYDROMETALLURGY OF SILVER
blende is present. Ores heavily charged with sulphides are not
suitable, especially if they contain a large percentage of galena
and zinc blende. These minerals need a low, gradually increas-
ing temperature, which conditions cannot be maintained in the
Stetefeldt furnace. If highly sulphureted ore is fed into the
shaft the temperature in the upper part of the shaft extend-
ing closely to the feeding machine becomes intense, and the
conditions required for roasting such ores become reversed.
Instead of being exposed to a low and then gradually increasing
temperature, the ore encounters the hottest zone first, and com-
bustion is so rapid that when it drops to the bottom it will be found
to consist mostly of minute globules, formed by partial melting
of the ore particles. Besides, much crust is formed at the bottom
of the shaft as well as on the sides, which keeps falling down in
large chunks. Even the iron grate on which the punched screen
of the feeding machine rests becomes clogged with incrustation,
which occurs so frequently as to make a regular feeding impos-
sible. This was the reason that while a large number of this
type of furnace were in successful operation on the Pacific slope,
in Nevada, California, Utah, etc., all the numerous attempts to
work the more mineralized ores of Mexico were failures.
For the proper ore, this furnace is undoubtedly the cheapest
to roast ores, especially for those which, on account of their
small percentage of sulphur, cause a large consumption of fiiel.
The conditions necessary to heat such ores quickly and effectively
are excellent, because every particle is exposed to the heat when
it passes in a shower through the flame.
G. Kiistel had once to roast in a Stetefeldt furnace ores which
were too poor in sulphur to produce sufficient chlorine. Sul-
phureted ore to mix with the dry ore could not be procured,
either. Mr. Kiistel solved the problem by burning brimstone in
a cast-iron pan and conveying the gas into the furnace near the
bottom of the shaft. This could not have been done successfully
with any other furnace.
IX
COLLECTING THE FLUE-DUST
THE formation of dust in roasting is unavoidable. In some
of the furnaces the amount of dust carried off by the draft is
comparatively small, as in the reverberatory and Bruckner,
while in others, like the White-Howell and the Stetefeldt, it is
excessive. The dust consists of ore particles, more or less roasted,
carried out of the furnace by the draft, and of condensed fumes
of volatilized metal chlorides and oxides. The former are much
easier collected than the latter.
The problem of collecting this dust is quite a difficult one, but
is very interesting and of great importance in metallurgy. The
idea of collecting the dust by conveying the gases into large
chambers in order to reduce the speed of their movement, thus
giving time for the dust to settle, is not a correct one. Close
observations have shown that in such a chamber a much larger
accumulation of dust will be found close to the walls than on
other parts of the floor. Besides, the walls will be found to be
covered with scale-like formations of the dust, the pointed
part of these scales turned against the current of the gases. When
these scales become too heavy they peel off and drop to the floor,
hence the larger accumulation of dust on the floor near the foot
of the walls. The cause of this greater precipitation can be no
other than the friction between the dust-charged gases and the
walls. It can also be observed that wherever the flue makes
a sharp turn the precipitation of dust is greater, because the
friction between wall and gases is much greater if the latter are
forced to make a sudden change in their course than if they are
permitted to follow a straight-line course. Furthermore, it can
be observed that, if an obstacle is placed in the flue, against
which the moving gases have to strike, it will cause a precipita-
tion of dust which increases with the swiftness of the gas current.
87
88
HYDROMETALLURGY OF SILVER
These observations demonstrate that a slow movement of
the gases, as attained with large dust-chambers, is by far not
so effective as a swift movement with increased friction, and
therefore much better results will be obtained if the dust-chambers
are so constructed as to offer a large wall surface as compared
with the area of the cross-section of the chambers, and if the
same are arranged in a zigzag fashion.
Based on the above observations I have devised and con-
structed a dust-collecting arrangement which is very effective and
gives much satisfaction.
14-8fc- >j
L-
FIG. 26.— VERTICAL SECTION OF HOFMANN DUST COLLECTOR,
0. Hofmanri's Flue-Dust Collector. — Fig. 26 represents in a ver-
tical section the dust-collecting arrangement in connection with
a White-Howell roasting furnace. A, feed end of the furnace;
J5, first chamber, the end of the furnace projecting a few inches
into it. Here the coarsest part of the dust accumulates and is
removed from time to time through the door H. From here the
gases pass through the opening, K, into the second chamber, C,
and from there through the arch, M, into the collecting shaft, D,
which they ascend, leaving it through the flue, F, and entering
COLLECTING THE FLUE-DUST
89
the second collecting shaft, E, in which they descend, leaving
the same through G, which makes connection with the general
flue leading to the chimney.
FIG. 27. —DETAILS OF BARS AND BEARINGS, HOFMANN
DUST COLLECTOR.
In the collecting shafts, cast-iron double channel irons are
arranged in rows, leaving a space 3 in. wide between each two.
These channels are closed at each end (Fig. 27) and extend as
round bars If in. in diameter and 1 ft. 8| in. long at the front
and 3 in. long at the opposite side. These two cylindrical exten-
4 — ,
Section, on Line A?B
T
1 /-:-;•- 1
?~T
' '
f 1*
I v-y i
i
[ft
Scale l^*=
Bearing for Bar Scale 1J$ - 1 ft.
FIG. 28. — DETAILS OF BARS AND BEARINGS,
HOFMANN DUST COLLECTOR.
sions rest in cast-iron bearings, as shown in Fig. 28. The longer
bar extends through the front wall of the shaft, and is cast square
at the very end, which part projects out of the wall, while the
shorter rests in a bearing inserted in the back wall of the shaft.
These channels can be turned by a socket wrench slipped over
90
HYDROMETALLURGY OF SILVER
the square end. Each row is 10 in. above the other, and the
position of the channels is such that the channels of one row are
placed right above the open spaces of the row below, as shown in
Fig. 29. The bearings are 1J in. in diameter while the cylindri-
cal parts of the bars are only If in., in order to allow room
for expansion and to permit an easy turning of the bars when
hot.
The gases from the furnace entering the shaft D, through M
(Fig. 26), in ascending partly strike the channels of the first row
and partly pass through the open spaces. That part, however,
which strikes the channel rebounds and is also forced to pass
through the open spaces. Passing through, they strike against
M- 8*-
i^. 29. — POSITION OF BARS, HOFMANN DUST COLLECTOR.
Bearing should be made If in. diameter, or J in. larger than the bar, so that the
latter can be easily turned when hot.
the channels of the next row above, rebound, and force their way
again through the open spaces. This is repeated until the gases
leave the top row and enter the second shaft, E. Here the same
play of the gases takes place, only that they have to descend
through the shaft.
It is apparent that the numerous objects placed in the path
of the gases and against which they have to strike, and the
large surface which they offer for friction will produce a very
effective precipitation of the fumes and dust. This is actually
the case. I erected at the works of the United Zinc and Chemical
Company, Argentine, Kansas, a system of three such collecting
shafts. Blende-pyrite ore was roasted in mechanical furnaces to
COLLECTING THE FLUE-DUST 91
produce sulphur dioxide gas for the manufacture of sulphuric
acid, and it was of importance that the gases should enter the
Gay-Lussac tower as free of dust as possible. After the system
was in operation for some time an investigation was made as to
its efficiency. The fumes passed first through a down flue and a
piece of straight flue, in which the coarse dust accumulated before
it entered the first shaft. It was found that the dust on the
bottom of the first shaft had the color of the roasted ore, showing
that the main part of it consisted principally of mechanical ore-
dust. The dust on the bottom of the second shaft was of a
light pink color, showing that the main part of it consisted of
precipitated white fumes of lead, iron and zinc. On the bottom
of the third shaft the dust was perfectly white, showing that all
the ore-dust was precipitated before it .reached the third shaft,
which was proved by actual analysis of this dust, showing it to
consist of the sulphates of lead, zinc and principally iron, while
almost no insolubles were present. In the first shaft were found
6.65 cu. ft. of dust; in the second, 4.15 cu. ft.,- and in the third
only 1.55 cu. ft. This very rapidly decreasing volume of dust
found in each succeeding shaft illustrates the great efficiency of
this system. I may mention as a further illustration that, at
the bottom of the down flue, where the coarse settled, the layer
of dust was 7 in. deep. In the piece of straight flue next to the
first shaft the layer was only 1J in. thick, while in the first shaft
it was 8 in.
The channels have to be shaken and turned from time to time,
depending on the dusting qualities of the ore and furnace. Once
or twice a week will be found sufficient. The shaking and turning
should, of course, be commenced at the top row. Fig. 30 shows
the dust-collecting and flue arrangement in connection with two
White-Howell furnaces.
This dust-collecting arrangement is compact, and very effect-
ive, and ought to be inserted in all the works where ores are
roasted, either in order to regain the valuable dust and the vola-
tilized silver, copper, lead, etc., or to prevent the entering of the
dust and volatilized substances into the subsequent chemical
process.
A very effective method of collecting the dust is the bag
system, in which the gases are forced by fans through long bags
made of burlap, flannel or muslin, which act as filters. The gases,
92
HYDROMETALLURGY OF SILVER
r1
COLLECTING THE FLUE-DUST 93
however, have to be first cooled sufficiently so as not to ignite
the bags, which cannot always be easily and cheaply accom-
plished. Besides, if, as in a sulphuric acid factory, the furnace
gases have to enter the process hot, the bag system cannot be
applied. This system is often used in smelting works.
. x
SULPHATING ROASTING
THIS mode of roasting, which has the object of converting
the silver into silver sulphate, in which state it is soluble in water,
is only used if silver is to be extracted with hot water by
Ziervogel's method.
The material to be suitable for this roasting has to consist
principally of copper and iron sulphides, of which the former has
to predominate, and has to be free of, or to contain only in small
quantities, the sulphides of lead, zinc, arsenic, and antimony.
For this reason it is exclusively used for argentiferous copper
matte. In this roasting the copper and iron have to be converted
into oxides, while the silver has to be changed into a sulphate.
The transformation of the silver into sulphate is done almost exclu-
sively by the sulphuric acid fumes which result from the
decomposition of cupric sulphate at a higher heat. Cupric sul-
phate and ferrous sulphate are formed in the first stage of roasting.
Ferrous sulphate is decomposed at a much lower temperature
than cupric sulphate, in fact at a temperature not high enough
for the formation of silver sulphate, so that the sulphuric acid
generated by the decomposition of the ferrous sulphate is of
very little avail for the formation of silver sulphate; it is, however,
of great effect in the formation of cupric sulphate, which then, at
a higher heat, sulphatizes the silver.
A certain percentage of iron sulphide is therefore advanta-
geous for this roasting process, but if the iron sulphide is in
excess the formation of silver sulphate, and with it the extraction,
suffers.
At Mansfeld, Germany, where this mode of roasting and the
subsequent extraction of the silver with hot water was originated
by Mr. Ziervogel, the roasting charge consisted of sulphur 19.32
per cent., copper 58, iron 9.18, lead 2.48, zinc 4,31, manganese
94
SULPHATING ROASTING 95
0.15, nickel 0.43, cobalt 0.83, silver 0.286, insoluble 1.08 per
cent., and permitted an extraction of 91 per cent, of the silver,
while at Schemnitz, Hungary, a matte containing 88 per cent,
iron sulphide and only 1.5 per cent., of copper sulphide, which
was tried by this method, permitted only an extraction of 73 to
75 per cent, of the silver.
This roasting is a very delicate process and has to be con-
ducted with great care and skill, otherwise inferior results will be
obtained.
At Mansfeld the roasting is done in a two-story reverberatory
furnace. The operations are as follows:
Six hundred pounds of pulverized copper matte are charged
on the upper hearth, spread, and about 5 Ib. of slacked bituminous
coal scattered over the charge, and stirred. This addition of
coal is made merely to help heat the charge in order to hasten
the operation. A matte richer in iron sulphide does not need
the addition of coal, because iron sulphide ignites quicker than
copper sulphide. The charge is stirred continually, and the
lumps which form have to be mashed with the furnace shovel.
They are soft and easily mashed, and are more numerous if the
material contains more iron than if it is poorer in iron. They
are caused by the conversion of the ferrous sulphate into basic
ferric sulphate, which melts easily. In this period the iron
oxidizes before the copper, and by the action of the sulphuric
acid changes into ferrous sulphate, which later at an increased
heat gives off sulphuric acid fumes and changes into ferric oxide
and basic ferric sulphate. The copper sulphide is converted into
cupric sulphate, but more by the acid fumes of the ferrous salt
than by the action of the air.
The time of roasting on the upper hearth is governed by the
time required to finish the charge on the lower hearth. During
this period, which lasts from five and one-half to six hours, the
charge has to be turned twice so that all parts of it are exposed to
the same heat.
When the lower hearth is clear, 25 Ib. of slack bituminous
coal is spread over the charge, which then is drawn to the drop-
hole in the bottom of the hearth, through which it falls to the
lower hearth. When the coal is mixed with the charge, burning
gases are emanating from the ore.
At the end of the first half-hour the ore on the lower hearth
96 HYDROMETALLURGY OF SILVER
commences to glow brighter than it did on the upper hearth,
caused by the higher temperature kept here and the further
oxidation of the sulphur. The thickness of the charge, which is
about 2J in., swells, on account of the burning sulphur, to 3£ or
4 in. In order not to burn the coal in the charge too quickly by
the action of the air, and to give it better opportunity to act on
the salts in the ore, the draft is very much checked while the
charge is raked continually and very briskly, in order to avoid
as much as possible the formation of lumps. This is done for an
hour, after which time all the coal is consumed. Then the charge
is turned, the part from the hotter place to the cooler, and that
from the cooler to the hotter place. After this the draft is in-
creased to its full capacity in order to produce a strong oxidation
by the inflowing air. This cools the charge after a while until it
becomes almost dark. To judge the end of this period, a sample
is taken from the middle of the hearth, cooled, the fine separated
from the lumps, and by means of a spatula a ridge is made of
the fine in a porcelain saucer. The saucer is held slightly inclined,
and some water, drop by drop, is poured behind the ridge. The
water is first absorbed by the sample, but after being saturated
a clear liquor slowly flows out from the other side of the ridge.
By the color of this liquor and its behavior toward salt the pro-
gress of the roasting is judged. If the roasting was conducted
right, by this time the liquor should have a clear blue color and
by the addition of some salt should give a light precipitate of
silver chloride, which is a sign that the silver sulphating has
commenced. If the liquor has a dirty greenish color it shows
that some ferrous sulphate is still undecomposed and a continua-
tion of the oxidizing period is required.
The coal, which is added to the charge and vigorously raked
and mixed with the latter while the draft is much checked, acts
on the neutral sulphates, which are converted into basic sulphates
while sulphurous acid escapes. . After the dampers are opened
and the full draft is given to the furnace, the sulphides which may
still exist will be completely roasted and all ferrous sulphate will
disappear, which is necessary to be accomplished before the sul-
phating of the silver takes place. The cuprous oxide oxidizes
to cupric, and at the end of this period the material should con-
sist of the free oxides of iron and copper, basic salts of iron,
copper and zinc, and neutral sulphates of copper, zinc, man-
SULPHATING ROASTING 97
ganese, some silver oxide and nearly all the balance of the silver
as sulphide.
When it has been ascertained by the above test that the roast-
ing has advanced to the proper stage, the ore is ready for the
sulphating of the silver. It will be remembered that at the end
of the previous period the material had cooled down almost to
darkness by the increased draft. The temperature has to be
increased again, but care is to be taken that this is done with a
clean oxidizing flame, and that the latter does not touch the ore,
so that none of the cupric oxide is reduced to cuprous oxide or
to metallic copper, as both of them would precipitate metallic
silver during the subsequent lixiviation, which silver would re-
main in the residues. For this purpose" very dry, thin limb-wood
should be used only. Pine is not to be recommended on account
of its pitch, which causes a smoky flame. The charge has to be
continually raked. After an hour it becomes dark red hot,
and later increases to bright red. The strong fire has to be kept
up uninterruptedly. If after two and a half hours the material
near the fire-bridge is completely roasted, the charge is turned
and roasting continued for one-half to three-quarters of an hour.
The material is properly roasted when the solution emerging from
the ridge of a sample on the saucer is only of a faint but clear
blue color, showing that but little cupric sulphate is present,
while by an addition of salt a heavy white precipitate of silver
chloride is formed.
If the heat is too high, some of the silver sulphate will be
reduced to metallic silver and all the cupric sulphate will be de-
composed, and therefore the liquor of the test will be colorless.
By increasing the heat to bright red during this period, the
fuming sulphuric acid liberated from the neutral cupric sulphate
does not act so energetically on the silver sulphide as the sulphuric
acid fumes resulting from the decomposition of the basic cupric
sulphate; hence the addition of fine coal at the beginning of
this period.
The time required for roasting a charge is eleven to twelve
hours, of which five and a half to six are consumed on the lower,
and, naturally, just as many hours on the upper hearth. All in
all, about four charges or 2400 Ib. of material will be roasted by
each furnace during twenty-four hours.
The loss of silver in Mansfeld was found to be 7.06 per cent.
98 HYDROMETALLURGY OF SILVER
which was caused partly mechanically by flue-dust, partly by
volatilization of silver oxide, which, however, in the cooler
regions of the dust-chambers decomposed into silver and oxygen.
In the roasted matte 91.74 per cent, of the silver was converted
into silver sulphate and was extractable, while 1.20 per cent,
silver remained in the residues.
XI
CHLORIDIZING OF ARGENTIFEROUS ZINC-
LEAD ORE
IN this chapter the detail records are given of investigations
of chloridizing roasting of argentiferous zinc blende and galena
ore, which I had the opportunity to make on a large working
scale.
The chloridizing roasting of this class of ore had not pre-
viously been made the subject of a thorough investigation on a
large scale, and the record of such experiments and investiga-
tions may be of interest and practical value.
The San Francisco del Oro mine is situated near Santa
Barbara and Parral, Chihuahua, Mexico. The ore contains on
an average 26 to 30 oz. silver per ton, besides a trace of gold.
The principal silver-bearing minerals are zinc blende, of which
the ore carries 37 per cent, and more, and galena, of which it con-
tains from 13 to 19J per cent. The heavy solid occurrence of
the ore and the great width of the vein permits very cheap
mining. The cost per ton does not exceed $1.50 Mexican cur-
rency, including hoisting and a slight hand-assorting.
(a) Zinc Blende. — The fine-grained black zinc blende pre-
dominates, containing about 25 oz. silver per ton; but there
occurs also a brown blende of a peculiar luster, somewhat resem-
bling bronze-colored mica. It is richer in silver than the black
blende, assaying from 55 to 70 oz. silver. The blende contains
considerable cadmium.
(b) Galena, with 40 to 50 oz. silver per ton, is finely im-
pregnated in the zinc blende, and can scarcely be detected with
the eye, and only a very small portion occurs as defined galena,
which makes it impracticable to lessen the percentage of lead
in the ore by hand-sorting. An attempt was made to sort out
the lead ore for shipping, but it was soon abandoned because the
99
100
HYDROMETALLURGY OF SILVER
amount of pure lead ore thus obtained was too small to pay for
such close work. Besides, the percentage of lead in the ore was
reduced only 0.5 per cent.
(c) Iron Pyrites. — Either intermixed with the zinc blende or
intersecting the same in narrow streaks. It contains about $12
gold per ton, but constitutes only a comparatively small per-
centage of the ore.
(d) Copper Pyrites. — Occurs seldom and then only in very
small quantities.
(e) Native Silver. — Now and then specimens with metallic
silver in flakes or wire are found.
(/) Gangue. — The minerals forming the gangue are quartz
and calcspar.
The ore looks like solid zinc blende; it is heavy and solid,
showing but very little gangue. In the following table two
analyses are given. Each one represents the average of large
lots of ore. The one is unassorted, just as it was extracted
from the mine, while from the other the galena and gangue were
sorted.
ANALYSIS OF SAN FRANCISCO DEL ORO ORE
UNASSORTED
ORE
ASSORTED
ORE
Zinc
2408
2550
Lead .
11 92
11.56
Iron
700
6.50
Manganese. . .
070
0.53
Cadmium
0.16
0.30
Antimony
0.50
0.52
CoDDer
0 72
1 02
YYrI~*
Alumina .
1 30
365
Calcium carbonate
982
8.00
Sulphur .... ...
21 35
21.01
Nickel
0.20
Silver
0.10
0.12
Gold
trace
trace
Soluble silica
092
Insoluble gangue
21.32
19.41
Total
i nn no
no 19
Near the surface the ore contained more free galena and less
zinc blende; and it is said that the first owners made quite a
financial success by smelting the ore in Mexican furnaces. How-
ever, when the character of the ore changed, the mine changed
CHLORIDIZING OF ARGENTIFEROUS ZINC^LEAD' 'ORE 101
hands, and all subsequent attempts to work the ore proved a
failure.. Though the ore was offered much cheaper than other
ore to the custom mills of Parral, only small quantities were
bought, being merely used as flux for the oxidized ores of the
Veta Grande and other mines of the district to facilitate chloridiz-
ing roasting. Many attempts, however, were made to work the
ore by itself, but without success. Even smelting, which ought
to have been out of the question, was tried. The main difficulty
was found to be in roasting. The ore caked very readily and the
silver could not be chloridized, at least not above 17 or 20 per
cent. After many unsuccessful attempts, further trials were
abandoned until an English company purchased the property.
The exceedingly refractory character of the ore and the very
discouraging experience of others induced the managing director
of the English company to have elaborate experiments made
before erecting a mill near the mine, the new mill to be constructed
in conformity with the observations and experience derived
from the experiments. For this purpose the Bosque mill at
Parral, an old 25-stamp lixiviating mill, was purchased, and I was
commissioned to conduct the experiments.
The Mill. — It consisted of 25 stamps, rock-breaker and self-
feeders, one large-size Stetefeldt furnace, claimed to roast 60
tons per day, one brick-lined revolving cylinder furnace of the
White-Howell type, 24 ft. long and 4 ft. in diameter, and two
leaching plants, one of 11 the other of 10 leaching vats, of vari-
ous sizes, averaging about 10 ft. diameter by 3 ft. 6 in. depth,
which, for want of grade, were almost buried in the ground.
The whole arrangement of the mill was ridiculously inconvenient,
causing a never-ending handling of the ore, supplies and products.
For experimental purposes, however, the mill was good enough,
especially as I had the privilege of erecting any other roasting
furnace which I should consider advisable to experiment with.
ROASTING EXPERIMENTS
Theory. — As shown by the analysis, the ore consisted prin-
cipally of the sulphides of zinc, lead and iron. The other sul-
phureted minerals occurred in such small quantities that it was
not necessary to pay any attention to the chemical part they
took in the process of chloridizing roasting.
102 HYDROMETALLURGY OF SILVER
Zinc blende, if subjected to the oxidizing action of the air,
is converted into zinc oxide and zinc sulphate, while sulphurous
acid escapes. In presence of salt, zinc sulphate remains indif-
ferent, and does not decompose the salt, at least not at the tem-
perature used in chloridizing roasting. If pure zinc blende, finely
pulverized, and mixed with salt, is placed on a roasting dish,
and roasted in the muffle, no chlorine gas can be detected, even
if exposed to a bright red heat. Even freshly prepared zinc
sulphate mixed with salt and exposed to the heat of the muffle
does not produce any chlorine gas. Zinc blende, therefore, does
not take an active part in producing chlorine during chloridizing
roasting, at least not enough to be of practical value. In the
roasted ore we find, therefore, the zinc mostly as oxide and sul-
phate and so also in the flue-dust.
If galena is subjected to a chloridizing roasting, especially in
presence of sufficient air, most of the lead is converted into a sul-
phate, which, as such, like the zinc sulphate, does not react on
the salt, and therefore does not generate chlorine.
Iron pyrites is converted into ferric oxide and into ferrous
and ferric sulphates, both of which react very energetically on
salt, and generate chlorine.
We have, therefore, two non-generators and only one generator
of chlorine in the ore. The non-generators of chlorine, galena
and zinc blende, however, contain all the silver, while the iron
pyrites carries only some gold, but no silver. This is a very im-
portant point to take into consideration. The next important
point is the fact that on account of the great density of the zinc
blende it requires a long time to oxidize. Likewise the galena
requires long roasting at a low heat, while iron pyrites decomposes
quickly, and in presence of salt generates chlorine at a period of
the roasting when neither the zinc blende nor the galena are
sufficiently oxidized to yield their silver to the action of the
chlorine. If, therefore, the salt is mixed with the ore in the
stamp-battery, the chlorine produced by the reaction of ferric
sulphate and salt is lost, and a very imperfect chlorination of the
silver takes place, no matter how long roasting may be continued
and how much salt may be used. For instance, when the ore
was roasted with 12 per cent, of salt in the Stetefeldt furnace,
the roasted ore contained but 1.38 per cent, of chlorine, of which
0.8 per cent, was combined with sodium, which represents 1.3 per
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 103
cent, undecomposed sodium chloride. The salt used in roasting
contained but 78.25 per cent, of sodium chloride, and the 12
per cent, of salt represents, therefore, 9.39 per cent. Not taking
into consideration the loss in weight which the ore sustained in
roasting, we can assume that 8.09 per cent, sodium chloride was
decomposed or volatilized, while not more than 15 per cent, of
the silver was chloridized, showing that the silver was not yet in
a proper condition to be acted upon by the chlorine at the time
when the reaction took place between the iron sulphate and the
salt. Only a very small percentage of chlorine was found
to have combined with other bases. The chlorine was prac-
tically an entire loss. Similar observations were made by
roasting in other furnaces.
In roasting an ore like the San Francisco del Oro ore, it is
therefore of the greatest importance to add the salt afterward
and not to mix it with the ore in the stamp-battery. But this
is not the only condition to be observed. We have to take into
consideration that in this case we are relying on the sulphates of
iron to generate chlorine, and that these sulphates easily decom-
pose, forming oxides. If, therefore, the oxidizing period is con-
tinued until the zinc blende and galena are well oxidized, which
takes a long time, we will have no iron sulphate left to decompose
the salt, and in consequence will have a very badly chloridized
ore. At a high temperature these iron salts decompose more
quickly, giving off their sulphuric acid, and in order to retain
them as long as possible the ore has to be roasted at a low tem-
perature. To know the proper time when the salt is to be added
is of the greatest importance; this knowledge, however, can be
obtained only by repeated tests and very close observation. The
most suitable time to add the salt for the San Francisco del Oro
ore was found to be when the black color of the ore turns brown
but still shows black particles. If at that time the salt is added,
a distinct odor of chlorine can be observed, which lasts during the
whole of the finishing period ; while, if the salt is added too soon
or too late, no chlorine evolves from the ore during this period.
The best results with this ore could undoubtedly be obtained
by subjecting it first to a dead oxidizing roasting, then adding a
mixture of green vitriol and salt. But the ore is not rich enough
to permit such an expense.
While the incapacity of the zinc and lead sulphates to de-
104 HYDROMETALLURGY OF SILVER
compose the salt makes the process of roasting difficult and
complicated, it offers, on the other hand, the advantage that it
reduces materially the consumption of salt. An addition of four
or five per cent, of salt gives the same result as eight or ten per
cent. If the roasting is very carefully conducted even three per
cent, gives good results.
Remarks. — The most difficult and at the same time the most
important process in the treatment of ores by wet methods is
undoubtedly the chloridizing roasting. It is always the safest
plan for the operator to roast as thoroughly as possible. If the
silver is well chloridized, the sodium hyposulphite will extract all
the silver chloride and frequently will leave the tailings even
poorer than indicated by the chlorination test, without the use
of additional solutions or chemicals, thus saving time and expense,
and not complicating the process. A high chlorination does not
necessarily involve a high loss by volatilization. I have
recorded instances in which the volatilization of silver was greater
in imperfectly chloridized charges than in well-chloridized ores
(see Chapter IV, "Loss of Silver by Volatilization"). Solutions, by
which we can correct a badly roasted charge, like chloride of
copper applied during base-metal leaching, or by which part of
the unchloridized silver can be extracted, like potassium cyanide,
or, as in some instances, Russell's extra solution, are very useful
and acceptable; but to neglect roasting and to rely for closer
extraction on these solutions is a rather dangerous practice.
Being convinced that the successful working of the San
Francisco del Oro ore depended on a successful roasting, and
knowing the great difficulty which the nature of this ore offered
to chloridizing roasting, particular attention was paid to this pro-
cess, and careful studies were made of the peculiarities of the ore.
The principal points to ascertain were: first, the mode of treat-
ment which the ore required with regard to temperature, roasting
time, draft, and the proper time for adding the salt; and, second,
which of the furnaces would comply best with the requirements
and at the same time perform the work the cheapest.
ROASTING IN THE STETEFELDT FURNACE
The furnace was a large-size Stetefeldt furnace, claimed to roast
60 tons per day, and was built according to the best improved design.
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 105
Though quite elaborate experiments were made, it was not
possible to obtain good results; in fact, they were far from being
satisfactory. Nevertheless they are interesting and, at the same
time, useful, inasmuch as they establish the fact that an ore like
the Del Oro, containing 25.5 per cent, zinc, 11.56 per cent, lead,
and 21 per cent, sulphur, is by far too refractory for the Stetefeldt
furnace. Such an ore requires to be submitted to a long and
gradually increasing temperature before the salt is added. The
principle of the Stetefeldt furnace, however, is just the reverse of
this important condition, and the results, as a matter of course,
could not be satisfactory.
Notwithstanding that there was a very powerful draft through
the furnace, and that the ore was crushed through 40-mesh
screen, only about three-eighths of the ore (by volume) was carried
by the draft into the descending flue and deposited at the bottom
of the chambers, while about five-eighths of it came down the
shaft. Actual dust was not carried beyond the second dust-
chamber, and even there only small quantities deposited, owing
to the great specific gravity of the ore.
After heating the furnace gradually for three days, charging
was begun. There were 20 stamps running, which crushed
through 40-mesh screen from 20 to 22 tons per twenty-four hours,
according to the amount of salt used. The experiments were
begun with an addition of 8 per cent, salt, increasing the amount
during the time of the experiments to 12 and finally to 16 per
cent. It was also tried to roast oxidizingly and to add the salt at
intervals to the ore at the bottom of the shaft, but without success.
The furnace was kept running for seven days and was then stopped,
as no signs of improvement in the work could be noticed and the
cooling floor was filled with badly roasted ore.
Before making any comments, the average results obtained
under different conditions are given:
ROASTING WITH 8 PER CENT. SALT
SHAFT
Oz. PER TON
DESCENDING FLUE
Oz. PER TON
Average of .raw ore including salt ....
Average of roasted ore .
32.37
34 41
32.37
2967
Average of leach tailings
28 62
23.72
Average of chlorination
16 90 per cent
20 20 per cent.
106
HYDROMETALLURGY OF SILVER
ROASTING WITH 12 PER CENT. SALT
SHAFT
Oz. PER TON
DESCENDING FLUE
Oz. PER TON
Average of raw ore including salt. . . .
Average of roasted ore
Average of leach tailings
Average of chlorination
31.00
31.33
26.60
15.20 per cent.
31.00
27.50
10.78
60.80 per cent.
ROASTING WITH 16 PER CENT. SALT
SHAFT
Oz. PER TON
DESCENDING FLUE
Oz. PER TON
Average of raw ore including salt ....
Average of roasted ore
30.32
33 39
30.32
2981
Average of leach tailings
28.35
22.30
Average of chlorination
15.20 per cent.
25.20 per cent.
An experiment was also made merely to oxidize the ore
without adding any salt:
OXIDIZING ROASTING
SHAFT
Oz. PER TON
DESCENDING FLUE
Oz. PER TON
Average of roasted ore
3491
28.57
Average of leach tailings
32.65
17.41
Average of extractable silver
6. 50 per cent.
39.20 per cent.
The oxidized ore from the descending flue was charged into a
tank, treated at first with a diluted solution of cupric chloride,
then leached with water and afterward with sodium hyposul-
phite, by which tailings were obtained containing 14.43 oz. per
ton, showing an extraction of 49.5 per cent, silver.
REROASTING THE ORE FROM THE SHAFT
The ore which had passed through the Stetefeldt furnace once
was sifted, to free it from lumps, and charged a second time.
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 107
SHAFT
Oz. PER TON
DESCENDING FLUE
Oz. PER TON
Average of roasted Ore
31 49
31 49
Average of leach tailings
Average of chlorination . .
29.60
9.40 per cent.
24.79
22. 30 per cent.
Observations and Comments. — If we compare the above
results we find that those obtained in the shaft were pretty
nearly equally bad, whether more or less salt was used. In the
descending flue 12 per cent, of salt gave the best result (60.8 per
cent, chlorination). An excess of salt lowered the chlorination.
The same observation was made afterward in the Howell and
reverberatory furnaces.
The ore from the shaft has a very dark, almost black, color,
and emits volumes of sulphurous acid gas when discharged, but no
chlorine. By leaving the ore in a pile on the cooling floor it
continues to emit sulphurous acid gas for a couple of days, without
a marked increase in the percentage of chloridized silver taking
place. In dropping through the shaft the main portion of the
ore is transformed into minute globules, which show that, while
the ore falls, it is partially slagged. Trying to avoid this, the
fire was lowered so much that the lower part of the shaft was
quite dark, while on the auxiliary grate the fire was allowed to go
out entirely. This, however, did not produce any change; the
ore came down now as before in globules, while in the descending
flue the temperature continued to be very high. It is apparent
that the ore when sifted into the shaft creates by the sudden
combustion of the sulphurets an extremely high temperature in
the upper regions of the shaft, which causes the suspended ore
particles to melt and slag to globules. Also that under such
circumstances silicates are formed, which, incrusting the ore
particles, prevent their further oxidation and chlorination.
An analysis showed that the roasted ore still contained 8 per
cent, of unoxidized sulphur. Ores containing not more than
8 per cent, sulphur usually roast well in a Stetefeldt furnace, and
it was expected that by charging the ore a second time good
results might be obtained. The ore was sifted, to free it from
lumps, and charged again. The results, however, proved to be
worse than those obtained in the first roasting. The chlorination
108 HYDROMETALLURGY OF SILVER
of 15 and 16 per cent, was reduced to 9.4 per cent. The ore
maintained its dark color and continued to emit heavy fumes of
sulphurous acid. It is not probable that part of the silver
chloride of the first roasting was decomposed by passing a second
time through the furnace; it is more likely that the dropping of
the chlorination in the shaft was caused by mechanical separa-
tion. Part of the lighter ore particles which contained less lead
and were better chloridized in the first roasting were carried
over into the flue during the second roasting, thus seemingly
reducing the original percentage of chlorination. Besides part of
the silver chloride may have slagged during the second roasting.
There are, however, other circumstances which also act dis-
advantageously. For instance, it was found after the first
roasting that the ore in the shaft contained 8.48 per cent, un-
oxidized sulphur and 11.19 per cent, lead, while that from the
descending flue contained only 0.51 per cent, unoxidized sulphur
and 3.11 per cent. lead. This shows that a separation takes
place, the shaft receiving the main portion of the lead, while the
lighter minerals, among them the iron pyrites, are carried over
into the descending flue. The better oxidation and chlorination,
as well as the higher temperature in the flue, are principally due
to this separation. But as the main bulk of the ore drops through
the shaft such a separation is disadvantageous.
Another interesting fact has to be recorded. Using so much
salt, and obtaining such an imperfect oxidation and chlorination,
we should naturally expect to find most of the salt undecomposed
in the ore. This, however, is not the case. In the shaft the
ore contained only 1.38 per cent, chlorine, of which 0.8 per cent.
was combined with sodium, representing only 1.30 per cent,
undecomposed sodium chloride. The material from the first
dust-chamber contained 0.38 per cent, of chlorine, of which 0.16
per cent, was combined with sodium representing only 0.27 per
cent, sodium chloride, while the material from the descending
flue contained but 0.2 per cent, chlorine. The fine white dust
from the last dust-chamber contained 0.42 per cent, chlorine and
a great deal of sulphuric acid. The salt, therefore, was decom-
posed, and the chlorine, either as hydrochloric acid or as chlorine,
escaped as gas without effect. If it were volatilized we ought
to have found more of it in the last dust-chamber.
Another bad feature is the formation of lumps. In the upper
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 109
region of the shaft, the ore-dust, wherever it comes in contact
with the hot walls, sticks to and incrusts them. This crust peels
off and drops down in pieces. It is almost raw, and some of the
larger pieces, when broken, show the texture of matte. They
form in large quantities. When the ore was sifted for reroasting,
there was not less than 25 per cent, of the whole ore in the shape
of hard lumps. These lumps contained only 21.87 oz. silver per
ton, and 0.82 per cent, chlorine, of which 0.28 per cent, was combined
with sodium, equal to 0.46 per cent, sodium chloride.
Besides the chemical difficulties, a very annoying mechanical
difficulty was encountered. The ore-dust in the same way as it
incrusted the walls of the shaft also incrusted the lower side of
the screen of the feeding machine, and thus stopped up the
holes, which necessitated a frequent changing of the screen, as
often as twice a day.
Heavily sulphureted ores, especially if they carry zinc, re-
quire more draft in roasting than less sulphureted ores. This is
especially the case with the Stetefeldt furnace, where the ore is
exposed for such a short time to the action of the air and
heat. In compliance with this theory all the air-doors with
which the furnace is provided were opened, but the result did not
improve.
REROASTING THE ORE OF THE STETEFELDT FURNACE IN
THE MODIFIED HOWELL FURNACE
The partly roasted ore from the Stetefeldt furnace, after
sifting, was fed into the Howell furnace. Having previously
ascertained that the ore contained only 1.3 per cent, of salt, 6
per cent, and sometimes 8 per cent, more salt were added. The
rate of feeding was changed from time to time in order to test
the working capacity of the cylinder for this ore; thus the rate
varied from 5 to 9 tons per twenty-four hours. The following
results are the averages of 33 charges:
Average of roasted ore 31.42 oz. silver per ton.
Average of leach tailings 17.55 oz. silver per ton.
Average of chlorination 44.20 per cent.
The consumption of wood in reroasting proved to be much
greater than in roasting the raw ore. The main portion of the
sulphur, especially that of the pyrites, which easily ignites, hav-
110 HYDROMETALLURGY OF SILVER
ing been burnt off in the Stetefeldt furnace, the ore did not create
any heat by itself, and all the required heat had to be furnished.
To reroast 8.3 tons it took 26 cargas of wood (12 cargas = 1 cord)
= 0.27 cord per ton, while it took only 16 cargas to roast 10 or 11
tons of raw ore, which is equivalent to 0.13 cord per ton of ore.
The roasting capacity of the furnace was not increased by roast-
ing this material, which contained only about 8 per cent, sulphur,
but on the contrary was diminished as compared with the raw ore
containing over 21 per cent, sulphur.
The reroasted ore was of a red-brown color, smelled of chlorine,
and did not emit any sulphurous acid gas, but it still consisted
principally of little globules. Quite a large portion of the globules
remained black, no matter how long the ore was kept in the fur-
nace. Some of them were magnetic, but the great majority
were not. Between the fingers the reroasted pulp felt sharp,
like pulverized glass. The temperature was kept at a proper
degree, and, the dust-chambers and furnace having been pre-
viously cleaned, there was an abundant draft, still it was not
possible to obtain more than 44.2 per cent, chlorination. The
cause of this failure wras undoubtedly the silicates which were
formed during the roasting in the Stetefeldt furnace.
APPLICATION OF STEAM
A jet of steam was introduced into the reverberatory of the
modified Howell. The hydrochloric acid which was formed by
the action of the steam was of decidedly beneficial influence, and
considerably improved the result; still the result did not give
entire satisfaction. The same percentage of salt was used and
the same temperature maintained as in the foregoing experi-
ments, and the improved results are therefore exclusively due to
the action of the hydrochloric acid on the silicates. The following
figures are the average of thirteen charges:
Average of reroasted ore 31.00 oz. per ton.
Average of leach tailings 10.37 oz. per ton.
Average of chlorination 66.60 per cent.
The ore still contained a considerable amount of these little
globules, but they had changed their color to red brown, and
between the fingers the ore felt soft and not so sharp and glassy
as when roasted without steam.
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 111
CONCLUSIONS
The experiments in roasting the argentiferous zinc blende and
galena ore of the San Francisco del Oro mine in a Stetefeldt fur-
nace have shown:
(1) An incomplete oxidation of the sulphureted minerals, the
main portion of the ore still containing 8.48 per cent, unoxidized
sulphur when roasted with salt, and 7.6 per cent, when roasted
without salt.
(2) An insufficient chlorination of the silver. The highest
chlorination in the shaft was only 16.9 per cent., and as 62.5 per
cent, of the whole volume of the ore dropped into the shaft, the
somewhat higher chlorination obtained in the descending flue
could not much improve the average chlorination.
(3) That the principle of the Stetefeldt furnace is contrary
to the conditions, of which the maintenance is so essential to
roasting ores containing much zinc blende and galena. Instead
of permitting the ore to be subjected for a longer time at a low
but gradually increasing temperature, the ore, entering the fur-
nace, encounters immediately the highest temperature, which is
detrimental to the roasting of such ores.
(4) That on account of the sudden exposure of the raw ore
particles to such a high temperature they melt to minute globules,
which makes the ore unfit for further treatment.
(5) That a concentration of the lead minerals takes place in
the shaft, which is disadvantageous.
(6) That about 25 per cent, of the ore when passing through
the furnace is changed into hard lumps of almost raw ore, and
that the construction of the furnace does not offer any means to
prevent it.
(7) That the lower side of the feeding screen becomes rapidly
incrusted and the holes obstructed, requiring a too frequent ex-
change of screens.
These observations taken together prove beyond doubt that
the Stetefeldt furnace is not suitable for the San Francisco del
Oro ore, and consequently for no ores heavily charged with zinc
blende and galena.
112 HYDROMETALLURGY OF SILVER
ROASTING IN THE WHITE-HOWELL FURNACE
The furnace which was used was not a regular Howell. It
consisted of a revolving cylinder of uniform diameter, the shell
being made of boiler iron and lined with bricks the whole length.
The principle on which it works, however, is identical with that
of the Howell, and this name is used here merely to indicate the
type of furnace.
The Howell, like the Stetefeldt furnace, requires the salt to
be added to the ore before entering the furnace. With some ores
this is immaterial, but it is a point of the greatest importance
for the Del Oro ore. If the salt is previously added the ore
becomes sticky, incrusts the furnace rapidly, and when it leaves
the furnace consists mostly of lumps, and what is still worse,
without being chloridized. If the ore is charged without salt it
remains dry and sandy, but a very annoying separation takes
place. The fine particles are carried by the draft into the dust-
chambers, and only the coarse sand passes through the furnace,
without being sufficiently desulphurized. If, then, salt is added
in the drop-pit, only a small percentage of the silver becomes
chloridized; the best results gave only 29 per cent, chlorination.
In order to diminish the separation, two per cent, of salt was
added to the ore in the battery, while the balance of the salt
was added in the drop-pit. This small percentage of salt made
the ore sticky enough to diminish considerably the dusting,
without causing the formation of lumps or too heavy an incrus-
tation of the furnace.
By this mode of roasting the chlorination improved consider-
ably, the average of three days' run being 67 per cent. It was
soon evident, however, that the Howell furnace, as such, could not
roast the Del Oro ore. The results were not uniform and relia-
ble, being sometimes high, sometimes low; and notwithstanding
the greatest care the average could not be brought above 67 per
cent. But the roasted ore was in a good condition; it was un-
finished but not spoiled, as in the Stetefeldt furnace, and there
was reason to expect that, by an alteration which would give
the ore more roasting time and allow a better regulation of the
temperature, good results would be obtainable.
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 113
ROASTING IN THE MODIFIED HOWELL FURNACE
In front of the furnace I constructed a shallow drop-pit and
a fireplace, the long side of the fireplace being opposite the dis-
charge of the furnace, so that the flame before entering the
furnace had to traverse the drop-pit. To one side of the drop-pit,
and communicating with it, there was attached a small reverber-
atory furnace 6 x 8 ft., the bottom of both being on the same
level. The reverberatory contained one working door and a
24-in. fireplace. When enough ore had accumulated in the pit
to make a charge for the reverberatory, it was pushed by means
of a hoe into the reverberatory. Each charge consisted of about
1400 Ib. While starting the furnace a strong fire was kept in
both fireplaces, but after the process was in operation the fire in
front of the cylinder was much lowered; in fact, so much so that
half the grate-bars remained bare of wood. Only now and then
a thin stick of wood was added, just enough to prevent the drop-
pit from getting chilled. Two per cent, of salt was added to the
ore in the batte^.
If the roasting is properly conducted, the blue flame of the
ignited pyrites can be observed in the back part of the cylinder.
Next to it and reaching beyond the middle of the cylinder the ore
assumes a higher temperature, forming a belt of bright-red heat.
In this region the principal oxidation takes place, and the increase
in temperature is caused by the oxidation and not by its position
nearer to the fire. The part of the furnace next to the fire and
nearly one-third of the whole length ought to look dark. The
furnace is mostly heated by the combustion of the sulphides,
and receives but little supply from the fireplace and from the
reverberatory. In fact, the ore while in the cylinder should be
left as much as possible to roast in its own heat. This is a very
important condition to maintain. The object is to convert as
much as possible of the galena and zinc blende into sulphates
and oxides before generating chlorine, and to avoid until then as
much as possible the decomposition of the iron salts. This can
only be done by maintaining a low heat after the combustion of
the pyrites. An excess of heat is invariably connected with an
excessive loss of silver by volatilization and by a low chlorination.
Galena and zinc blende roast quicker and better in a low than in
114 HYDROMETALLURGY OF SILVER
a high heat. When the ore leaves the cylinder and drops into
the pit it should be of a very dull red heat, while the color after
cooling should be dark yellow-brown.
If the temperature is so kept, neither the odor of chlorine nor
much of sulphurous acid can be detected. At an increased heat,
sulphurous acid emits again strongly, showing that the oxidation
is not yet completed. As the temperature in the cylinder is
mostly produced by the combustion of the sulphureted material,
the main means of regulating the same is the feed. If too much
ore enters the furnace the belt of bright-red heat increases, ad-
vancing more and more toward the front, and finally the whole
furnace assumes this temperature. The ore dropping into the
pit is very hot, emits heavy fumes, and overheats the pit. If
then removed into the reverberatory, it takes a very long time
to be finished, necessitating an interruption in the feed of the
cylinder. On the other hand, if insufficient ore is charged, the
belt of bright-red heat gets smaller and moves toward the back
end of the furnace.
When the properly prepared ore enters the reverberatory
furnace, the salt is added and the temperature is somewhat
increased. It commences to fume, and swells, without forming
more lumps than an ordinary ore. In the beginning strong
fumes of sulphurous acid emit, but soon cease, and chlorine
appears. The charge is finished if the fumes assume a mild and
sweetish smell of chlorine; as long as they smell strong, roasting
has to be continued.
A series of experiments was made to ascertain the smallest
amount of salt practicable, and it was found that 4, 6, 8, and 10 per
cent, give about equal results. Twelve per cent, begins to make
the ore too sticky and produces less chlorination; 3 per cent, is
sufficient if the roasting is very carefully conducted, but then
only 1 per cent, has to be added in the battery and 2 per cent, in
the furnace. Four per cent., however, is safer, as then the result
does not depend so much on the skill and good-will of the laborers.
The roasting capacity of the furnace proved to be much less
for this ore than for ordinary ore. Not more than 8J tons could
be roasted in twenty-four hours. It is true the cylinder was only
24 ft. long, but even with a 32-ft. cylinder it cannot be expected
to roast more than 12 tons per day. Each charge had to remain
two hours on the reverberatory hearth. Though the ore was
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 115
roasted in the reverberatory at a somewhat increased heat, yet
the temperature could not be increased beyond dull red without
losing too much silver by volatilization.
ADDITIONAL CHLORINATION AFTER THE ORE HAS LEFT THE
FURNACE
Some ores gain much in chlorination if left hot in a pile for
some time. This is mostly the case when an ore is insufficiently
roasted, or when the nature of the ore is such as to require a long
roasting time at a low heat. Another additional chlorination
can be produced by moistening the ore and leaving it for several
hours in a pile. This is usually the case if the ore contains copper.
Roasted ore containing caustic lime should not be left moist on
the cooling floor. The most important additional chlorination,
however, takes place, according to numerous observations of
mine, during base-metal leaching. The roasted Del Oro ore
either contained much caustic lime — which, however, is hardly
possible, as the raw ore is so rich in sulphur — or some other chemi-
cal substance which acted decomposingly on the silver chloride,
because it could not be moistened on the cooling floor without
sustaining quite a loss in chlorination. To prevent this, the vats
were charged about one-third with water and the hot ore dumped
into it, thus producing a hot base-metal solution. The observa-
tion was made that by this practice not only was the. decompo-
sition of the silver chloride avoided, but that a considerable
increase in the silver chlorination took place, in some instances
as much as 12.9 per cent, (see following table, charge No. 11).
If the original chlorination, however, was 75 per cent, or more,
this additional chlorination amounted to much less. By adding
some cupric chloride to the water in the vat before dumping the
ore, I found that badly roasted charges gained in chlorination as
much as 34 to 38 per cent, (see table, charges Nos. 9, 15 and 16).
These are very important observations, and give the operator
the means of correcting badly roasted charges.
RESULTS
The table on page 1 16 is a record of the results obtained in roast-
ing the San Francisco del Oro ore in the modified Howell furnace.
It represents a two weeks' run. As each tank charge contained
116
HYDROMETALLURGY OF SILVER
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CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 117
the whole ore of twenty-four hours' roasting, it offered a good
opportunity to follow each charge from the raw ore down to the
tailings, and to ascertain for each charge the loss by volatiliza-
tion, gain by additional chlorination, etc.
Taking the averages of the results, we find the silver chlori-
nation when the ore left the furnace, 68.4 per cent.; additional
chlorination, 13.3 per cent, or a total chlorination of 81.7 per
cent. The low average chlorination of the ore when leaving the
furnace was caused by the three badly roasted charges — 9, 15
and 16. The other eleven charges gave an average of about 75
per cent.
The total or actual chlorination of 81.7 per cent, may seem
to be low, but if we consider that the ore is of low grade, averag-
ing only 28.8 oz. per ton, and that 1 per cent, represents only
0.28 of an ounce silver; and also consider that the ore contains
about 37 per cent, zinc blende and 13 to 19£ per cent, galena,
which carry all the silver, and that the ore was pronounced as
being too refractory for chloridizing roasting, we have to count
the work done by the modified Howell furnace as very satisfac-
tory, especially as such a chlorination secured the success of the
enterprise owing to the cheapness of mining and reduction.
Loss OF SILVER BY VOLATILIZATION
The loss of silver by volatilization in these experiments was
ascertained by the method described in a previous chapter.
Owing to the fact that a great portion of the lead and zinc
sulphides is converted into sulphates, the San Francisco del Oro
ore loses but a small percentage of its weight during roasting.
The tests showed a loss of 2^ and 3J per cent.
With these figures, and the assay value of the raw and roasted
ore, the loss of silver by volatilization was calculated. The
extremes were 1.3 per cent, and 15.5 per cent, while the average
gave 7.9 per cent. The figures contained in the corresponding
column of the table illustrate how variable this loss is, and what
a severe loss of silver can be caused by even slight oversights.
I found the Del Oro ore to be more sensitively disposed for such
loss than many others, even antimonial ores which I had treated
before. The least increase of the temperature above dull red
causes a heavy loss, even if this increase of the temperature lasts
118 HYDROMETALLURGY OF SILVER
only a very short time. Thus two or three thin sticks of wood,
if thrown on the fire before needed, may materially increase the
loss.
The loss by volatilization is not in direct proportion to the
per cent, of chloridized silver. Frequently a well chloridized
ore suffers much less loss than a badly chloridized one. Refer-
ring to the table we find, for instance, in charge No. 16 the silver
was chloridized only to 47.2 per cent, while the loss by volatiliza-
tion was as high as 13 per cent. Again, in charge No. 20, 76 per
cent, of the silver was chloridized, while the loss by volatilization
amounted to only 1.3 per cent. We find the same in charges
Nos. 12, 14, etc.
THE ROASTED ORE
The roasted ore contains only a small percentage of lumps.
These are not hard, but porous, and fall to powder if kept in con-
tact with water for some time. If the ore is left dry in a pile it
hardens. If left undisturbed for a week or two it becomes so
hard that it requires the use of a pick to loosen it. In water,
however, it softens easily again. The color is usually red-brown,
but occasionally, if there is less iron pyrites in the ore, it is yellow-
brown.
The analysis of the unassorted ore after roasting (see above
analysis of the raw ore) is here given. The ore was roasted with
5 per cent, of salt.
ANALYSIS OF UNASSORTED ROASTED ORE
Gold trace.
Silver 0.09
Lead 9.00
Iron 6.00
Zinc 22.45
Lime (calculated as caustic) 5.65
Antimony 0.75
Copper 0.60
Cadmium 0.10
Alumina 3.09
Soda (calculated as caustic) 3.79
Sulphuric acid 13.16
Chlorine 0.88
Soluble silica 8.00
Insoluble gangue 18.61
Oxygen of the oxides ?
This analysis shows that the heavy metals were principally
converted into sulphates, and that only a small portion of them,
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 119
if any, can be present as chlorides. The 0.88 per cent, of
chlorine may be due to undecomposed salt.
If the furnace crust is not removed from the furnace for some
time, it changes its color. In some parts it is greenish white; in
others, flesh-colored. It gets very hard, and when moistened
with water generates heat and slacks like lime. I had never
made this observation before, not even with the very calcareous
ore of Las Yedras. It is to be regretted that it was overlooked
to make an analysis of this crust. It is difficult to believe that
this phenomenon could be caused by caustic lime, because the
latter could not well exist in an atmosphere of sulphuric and
sulphurous acid and of chlorine.
CONSUMPTION OF WOOD
Owing to the large quantity of sulphureted minerals in the
ore, and the very low temperature at which the Del Oro ore has
to be roasted, the consumption of wood is very small. After
the furnace is heated and the cylinder incrusted, it takes hardly
any fire in front of the cylinder to maintain the proper tempera-
ture. In the reverberatory addition a little more fire is needed,
but much less than ordinary ores require.
During two weeks the wood was weighed (it was bought by
weight) and the total consumption during this time was found to
be 220 cargas of 300 Ib. With this amount of wood 115.8 tons
of ore were roasted, which gives 1.8 cargas per ton of ore.
Twelve cargas of the Parral wood are equal to one cord, and
if we express the consumption in cords we find that with one
cord of wood the furnace roasted 6.3 tons of ore.
COST OF ROASTING IN THE MODIFIED HOWELL FURNACE
The cost of roasting 8J tons per twenty-four hours was as
follows :
Labor. . .................................... $6.60
4 per cent, salt, 680 Ib. at 1.27? ............... 8.63
15.7 cargas wood at 75^ ....................... 11.77
Steam power, 10 cargas wood at 75^ ............ 7.50
Oil, light, tools, etc ........................... 2.00
Management, office, mechanic's assay office ...... 1.78
8.5 = S4.50
Cost per ton ............................ $4.50 Mexican currency.
120 HYDROMETALLURGY OF SILVER
To ascertain the cost of steam power a separate boiler was
used for the furnace. It is apparent that by using a boiler for
only one small furnace the expense per ton of ore will be much
greater than if with the same boiler and engine several large
furnaces are operated. But this was the only way of getting an
estimate. The steam for working the pumps and preparing the
calcium sulphide was supplied by the same boiler, and had to be
charged to roasting.
As the statement of cost is made only 8J tons per day, it
would be misleading if the whole expenses for management,
mechanic's assay office, etc., should be charged to the 8J tons,
as those expenses will be about the same for 100 tons per day,
the intended capacity of the new mill. The expenses for manage-
ment, etc., were, therefore, calculated for 100 tons per day and
the 8J tons charged in proportion. But there are three depart-
ments in the mill, viz., stamping, roasting, and leaching, and
each department has to be charged with one-third of this expense.
The above item of $1.78 represents, therefore, one-third. In the
statement of cost 4 per cent, salt was put down because subse-
quent experiments proved this amount to be sufficient.
SUMMARY
It must be borne in mind that the figures contained in the
above table are the results of experiments. In other words, these
figures were obtained under different treatments with regard to
salt, temperature, time, etc., and the averages, therefore, do not
represent the best obtainable results. This is especially the case
with the loss by volatilization, which will be less after the men
acquire more skill in maintaining the proper temperature. But
as the roasting results obtained with the modified Howell furnace
under above conditions are good enough to secure a profitable
reduction of the ore, we may just as well accept these averages
as a basis for calculations and estimates.
RECAPITULATIONS
Average value of raw ore without salt 28.85 oz. per ton.
Average value of raw ore containing salt 27.49 oz. per ton.
Average value of roasted ore 26.10 oz. per ton.
Average value! of vat tailings 4.76 oz. per ton.
Average per cent, of chlorination 81.6 per cent.
Average per cent, of actual extraction 74.9 per cent.
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 121
Average number of ounces silver extracted per
ton 21.9 ounces.
Average percental loss by tailings 17.0 per cent.
Average percental loss by volatilization 7.9 per cent.
Average per cent, of salt used 4.7 per cent.
Average number of tons roasted per day with
one furnace 8.5 tons.
Average cost of roasting one ton of ore $4.50 Mexican currency.
Average consumption of wood per ton, includ-
ing steam power 3 cargas.
ROASTING IN THE REVERBERATORY FURNACE
The appliances for roasting consisted of a large size Stetefeldt
furnace and one 24-ft. revolving cylinder furnace, as described
above. After the Stetefeldt furnace proved to be a failure with
the Del Oro ore, the roasting capacity became reduced to that of
the revolving cylinder, or 8J- tons. In order to increase the
roasting capacity to the stamping capacity, and for the sake of
further experiments, four two-story reverberatory furnaces were
erected, the lower hearth of 220 sq. ft. surface and the upper of
210 sq. ft. surface (Figs. 10, 11, and 12).
Each two-story furnace took four charges of one ton each.
When one charge was finished all the others were moved forward,
and on the first hearth a new charge dropped through an opening
in the arch. From the second hearth the charge was dropped on
the lower hearth, through an opening in the bottom near the
working door. The upper hearth was used exclusively for oxi-
dizing roasting, and the 4 per cent, of salt for better mixing was
added while the charge was dropped on the lower hearth. As
the proper time when the salt is to be added had proved to be a
factor of great importance in roasting this ore, the process was
conducted on the appearance of the ore at the second hearth.
When the charge on that hearth showed it to be in proper condition
to receive the salt, the ore from the finishing hearth was dis-
charged, the charge from the third hearth moved on the finishing
hearth and the charge on the second hearth dropped on the
third, while the salt was added. Every two and one-half to three
hours a charge was done, and therefore each two-story furnace
roasted from 8 to nearly 10 tons per twenty-four hours, according
to the quantity of lead contained in the ore. Zinc blende roasts
quicker than galena. Each charge was ten to twelve hours in
the furnace.
The chlorination results were so near those obtained in the
122 HYDROMETALLURGY OF SILVER
modified Howell furnace that no details need to be given here,
but details and observations will be given of a charge which was
subjected to a prolonged oxidizing roasting, because they are
rather interesting. A charge of the Del Oro ore was placed directly
on the hearth nearest to the fireplace (finishing hearth) and kept
there until finished.
OXIDIZING ROASTING
First hour. — Assay of raw ore, 35.57 oz. per ton.
During this hour the ore had just fairly started to ignite. A
sample when cold had not changed its color, and looked like raw
ore. In this and the following hours a sample was leached first
with water, then with sodium hyposulphite, and to each filtrate
calcium sulphide was added.
Wash- water: no precipitate.
Hypo-solution (1 per cent.) : no precipitate, light coloration.
Concentrated hypo: a heavy precipitate, consisting mostly
of iron and lead.
Assay of ore, 33.54 oz. per ton; after leaching with hypo, 33.54
oz. per ton; no soluble silver.
Second hour. — Toward the end of this hour the ore com-
menced to lose its own heat, caused by the combustion of the
pyrites. During this hour no fire was kept up. The color of the
sample when cold was a dark greenish-brown.
Wash- water: no precipitate or discoloration, therefore no
salts soluble in water.
Hypo-solution (1 per cent.): a slight precipitate.
Hypo-solution (3 per cent.): a heavy precipitate.
Assay of ore, 34.41 oz. per ton; after leaching, 34.12 oz. per
ton; no soluble silver.
Third hour. — During this hour the ore had lost its own heat,
and fire was started again, but the temperature was kept very
low. The color of the sample when cold was brown-yellow,
more brown than yellow.
Wash- water: no precipitate, no coloration, no soluble salts.
Hypo-solution (1 per cent.): considerable precipitate.
Assay of ore, 36.74 oz. per ton; after leaching, 35.28 oz. per
ton; soluble silver, 1.46 oz. per ton, or 3.9 per cent.
Fourth hour. — The temperature was somewhat increased
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 123
during this hour, but still kept rather low. The color of the
sample when cold was of a much lighter yellowish brown.
Wash-water: the first indication of precipitate, showing that
until the end of the fourth hour no zinc sulphate had been formed.
Hypo-solution (1 per cent.): considerable precipitate.
Assay of ore, 35.57 oz. per ton; after leaching, 31.78 oz. per
ton; soluble silver, 3.79 oz. or 10.6 per cent.
Fifth hour. — The temperature still kept at dull red. Color
of sample when cold brown-red, showing that at this time some
oxide of iron had formed.
Wash- water: considerable precipitate of a yellowish-white
color, mostly zinc and cadmium.
Assay of ore, 35.13 oz. per ton; after leaching, 22.89 oz. per
ton; soluble silver, 12.24 oz. per ton, or 34.8 per cent.
Sixth hour. — The same low heat; color of sample when cold
darker red-brown.
Assay of ore, 33.83 oz. per ton; after leaching, 20.41 oz. per
ton; soluble silver, 13.42 oz. per ton, or 39.6 per cent.
Seventh hour. — The same temperature; sample when cold
still darker red-brown.
Assay of ore, 30.74 oz. per ton; after leaching, 13.41 oz. per
ton; soluble silver, 16.33 oz. per ton, or 54.9 per cent.
Eighth hour. — The same moderate roasting temperature.
No more sulphurous acid gas could be noticed. The ore com-
menced to look dead. The color of the sample when cold did not
change in this nor in the following hours.
Assay of ore, 30.62 oz. per ton; after leaching, 13.55 oz. per
ton; soluble silver, 17.05 oz. per ton, or 55.7 per cent.
Ninth hour. — The temperature was slightly increased, but
the ore did not fume, nor could sulphurous acid be noticed.
Wash-water: a very heavy, yellowish-white precipitate.
Hypo-solution (1 per cent.): a heavy precipitate, consisting
principally of lead, zinc and silver, no copper.
Assay of ore, 29.30 oz. per ton; after leaching, 12.09 oz. per
ton; soluble silver, 17.21 oz. per ton, or 58.8 per cent.
Tenth hour. — Temperature again slightly increased; no fumes,
no sulphurous acid noticeable. The ore remains fine and loose;
no lumps.
Assay of ore, 29.16 oz. per ton; after leaching, 12.24 oz. per
ton; soluble silver, 16.92 oz. per ton, or 58.0 per cent.
124 HYDROMETALLURGY OF SILVER
Eleventh hour. — During this hour the temperature was con-
siderably increased, bright red; the ore remained fine and loose;
no lumps; no sulphurous acid noticeable.
Wash-water; a yellowish-white precipitate of very clear color.
Assay of ore, 30.03 oz. per ton; after leaching, 13.26 oz. per
ton; soluble silver, 16.77 oz. per ton, or 55.8 per cent.
As the ore during this hour presented the appearance of a
dead roasted ore, oxidation was not carried any further, but 6
per cent, of salt was added at the end of the eleventh hour, to
make some further observations.
At the end of the eleventh hour and just after adding 6 per cent,
salt. — The salt was added, quickly stirred and a sample taken.
Wash-water: dark -colored precipitate.
Assay of ore, 28.74 + 6 per cent. = 30.46 oz. per ton; after leach-
ing, 11.66 + 6 per cent. = 12.36 oz. per ton; soluble silver, 18.80
oz., or 59.4 per cent.
Twelfth hour. — The temperature lowered for chloridizing.
The ore commenced to fume after the salt was added, but the
fumes were very thin and light.
Wash-water: black precipitate.
Assay of ore, 28.13 + 6 per cent. = 29.81 oz. per ton; after
leaching, 13.12 + 6 per cent. = 13.90 oz. per ton; soluble silver,
15.90 oz. per ton, or 53.3 per cent.
Thirteenth hour. — Same temperature; a mild, faint odor of
chlorine perceptible; no lumps, and when the ore was left in a
heap to cool it did not harden like other charges roasted with
salt. At the end of this hour the charge was removed from the
furnace.
Wash-water: dark precipitate, almost black.
Assay of ore, 28.42 + 6 per cent. = 30.12 oz. per ton; after
leaching, 11.07 + 6 per cent. = 11. 73 oz. per ton; soluble silver,
18.39 oz., or 61 per cent.
I found that the assay value of the ore during oxidizing
roasting dropped from 35.57 oz. per ton to 29.74 oz. This was
caused principally by an increase in weight which the ore sus-
tained by oxidation, especially the lead and zinc of which a large
percentage is converted into sulphate if roasted at such a low
heat. This took place principally during the seventh hour,
when the ore assumed a dark red-brown color. I further found
that during oxidizing roasting considerable of the silver is con-
CHLORIDIZING OF ARGENTIFEROUS ZINC-LEAD ORE 125
verted into a state in which it is soluble in sodium hyposulphite.
It shows first at the end of the third hour (3.9 per cent.) and gradu-
ally increases until it reaches its maximum (58.8 per cent.) at the
end of the ninth hour, and then diminishes during the next two
hours to 55.8 per cent.
By adding 6 per cent, of salt and continuing to roast for two
hours, the percentage of soluble silver increased only 2.2 per
cent, above the maximum obtained in oxidizing roasting, not-
withstanding that at the time the salt was added the ore was not
dead roasted, but still contained sulphates soluble in water.
The salt soluble in water, however, was zinc sulphate, which
does not act on the salt; the iron sulphate was decomposed by
that time. This experiment illustrates the great importance of
adding the salt at a certain time during oxidizing roasting —
that is, at a time before the iron sulphate is decomposed and the
oxidation of the lead and zinc sulphides has fairly advanced.
Only a part of the soluble silver in the oxidized ore was present
as a sulphate. From a sample of nine hours' oxidizing, when
leached with water, only 7.39 oz. silver per ton could be ex-
tracted, while the sodium hyposulphite extracted 17.21 oz. per ton.
In other words, of the 58.8 per cent, of soluble silver only 25.2
per cent, was sulphate of silver, while the balance of 33.6 per
cent, was some other silver salt, not soluble in water but soluble
in sodium hyposulphite, probably silver antimonate, though the
raw ore contained only one-half per cent, of antimony.
TREATING THE OXIDIZED ORE WITH CUPRIC CHLORIDE
A sample of the nine hours' oxidizing roasting was moistened
with a dilute solution of cupric chloride and left in that condition
for three hours, then leached with water and hypo solution:
Assay of ore, 29.30 oz. per ton; after leaching 7.29 oz. per ton;
soluble silver, 22.01 oz. or 75.2 per cent. Increased extraction
by using cupric chloride 4.8 oz. per ton, or 19.4 per cent.
CONSUMPTION OF WOOD IN THE REVERBERATORY FURNACE
The furnaces were built in pairs, two being connected with
one flue. During four weeks the wood consumed by one pair of
these furnaces was weighed. During this time 507 tons were
roasted at a consumption of 672 cargas of wood, or 1.3 cargas per
126 HYDROMETALLURGY OF SILVER
ton. If expressed in cords, we find that one cord of wood roasted
nine tons of ore, which is an exceedingly small consumption.
COST OF ROASTING IN THE REVERBERATORY FURNACE
Statement of cost for two two-story furnaces roasting 18 tons
per day:
Labor $22.60
23.4 cargas of wood at 75^ 17.55
4 per cent, salt, 1440 Ib. at 1.27^ , 18.28
Tools, etc 4.00
Management, office, mechanics, etc 3.77
$66\20-18=$3.67
Cost per ton $3.67 Mexican currency.
CONCLUSIONS
To form an opinion as to which roasting furnace is the most
suitable for the San Francisco del Oro ore, we have to take into
consideration only the modified Howell and the reverberatory.
The Stetefeldt did not roast the ore, and the Bruckner was not
tried, because I knew by experience that a large Bruckner furnace
could not roast more than five or six tons per day of such a
heavy ore, and therefore refrained from incurring the expense of
erecting a Bruckner furnace just for experimental purposes.
Both furnaces, the modified Howell and the reverberatory,
gave about the same results, and the loss by volatilization was
also nearly the same. The cost of roasting, however, is different.
The cost per ton of ore in the modified Howell is found to be $4.50,
and in the reverberatory $3.67. Difference in favor of reverbera-
tory, 83c. per ton. Besides this, the reverberatory furnace
creates much less flue-dust than the Howell. The latter forms a
great deal of flue-dust which is far from being roasted, even if
provided with an auxiliary fireplace. The labor question has
also to be taken into consideration. The reverberatory requires
more manual labor, and in a locality where labor is scarce, as is
often the case in Mexico, it may be more advantageous, under
certain conditions, to adopt the Howell notwithstanding the
greater cheapness of the reverberatory.
XII
CHLORIDIZING OF CALCAREOUS ORES
I WAS engaged by the Anglo-Mexican Mining Company to
investigate and improve the roasting of their ores at Yedras,
Sinaloa, Mexico.
The ore is treated by the lixiviation process with sodium
hyposulphite, and consists of argentiferous arsenical pyrites,
fine-grained black zinc blende, arsenical fahlerz, some iron py-
rites, and occasionally ruby silver, while the gangue consists of
silicious limestone and calcspar.
The chloridizing roasting of highly calcareous silver ores in
combination with argentiferous arsenical pyrites has, like the ore
treated in the preceding chapter, not often, if ever, been the
subject of a thorough investigation on a large scale, and in the
following pages the results and observation of such an investi-
gation are given in detail.
The only analysis of the ore which I could obtain is the fol-
lowing, made in San Francisco several years before the experi-
ments were made:
Per cent.
Silica 15.13
Sulphur 13.31
Arsenic 9.82
Iron 17.33
Alumina 1 .35
Zinc 4.92
Lead 1.78
Carbonate of lime 33.78
Magnesia 2.58
This analysis, however, does not represent the average of the
ore which was delivered from the mine to the mill. Frequent
concentration tests showed that the mill ore contained much
more gangue matters than the analysis shows. However, it may
serve to give a general idea of the ore. While experimenting I
127
128 HYDROMETALLURGY OF SILVER
felt very much the want of a chemical laboratory and a chemist
at the works.
The roasting facilities at the Yedras mill consisted of four
revolving Bruckner cylinder furnaces, each 16 ft. long, and eight
long reverberatory furnaces. The Briickners had been abandoned
for several years, because former operators, so I was told, could
not get satisfactory res'ults with them. The chlorination was
exceedingly low (40 to 50 per cent.), and the loss of silver by
volatilization extremely high, while the main part of the roasted
ore was rolled up into balls ranging from the size of an orange
up to 15 in. and upward in diameter, the inside of which was
not roasted. The subsequently erected reverberatory furnaces
also gave very poor results — 65 to 70 per cent, chlorination, with
a loss of silver by volatilization of 20 to 35 per cent, though
fewer lumps were formed.
ROASTING IN THE BRUCKNER FURNACES
I knew by experience that as a rule ores can be roasted with
less loss of silver by volatilization in the Bruckner than in the
reverberatory furnace, and as the enormous loss of silver which
the ore had sustained in the reverberatory was the most important
question, I started the long-abandoned Bruckner furnaces again.
The previous failures with them I ascribed to the application of
too high a temperature and an insufficient supply of air. The
formation of balls I expected to diminish by judiciously regulating
the heat and the revolving speed of the furnace, and if I succeeded
in chloridizing well these lumps, their formation would not be a
serious matter, for they could be pulverized in a ball-mill before
charging the ore into the leaching vats.
My improvements on the Bruckner cylinder, which consist of
a fireplace and flue arrangement attached to each end of the
furnace, enable me to apply the flame alternately through either
end, and thus to use much longer cylinders, and provide fully for
the free access of air between the fireplace and the throat of the
furnace, which, strange to say, is not the case with the common
Bruckner furnaces. In a reverberatory furnace air can enter
through the working doors; Stetefeldt provided his furnace with
air-channels through which the required supply can be regulated;
the Howell furnace is provided with an air-door; but the common
CHLORIDIZING OF CALCAREOUS ORES 129
Bruckner furnace has no proper means of regulating the supply
of air, and either the fire-door must remain open or only a very
limited amount of air can enter the furnace. If the whole fire-box
is built of brick it is easy enough to make the proper change, but if
it is made of boiler-iron lined with brick, it is not easy to make the
necessary alterations in a remote mining camp. In this case, the
fire-boxes being of brick, I added 12 in. to their length, moved
the grate bars toward the front, and inserted an air-channel
behind the new fire-bridge. The fresh air entering through this
chamber not only assisted in oxidizing the ore, but also aided
the combustion of the fuel. The speed of the furnace, which had
been 2J revolutions per minute, was reduced to one revolution
in If minutes, which materially reduced or altogether avoided the
formation of large balls and diminished the quantity of dust,
while on the other hand the speed was sufficient to expose every
particle of the ore to the action of air and heat; in fact, a still
slower speed, possibly even one revolution in three or four minutes,
would have been preferable had it been obtainable.
The ore was crushed dry in the stamp battery with 7 per cent,
of salt, and passed through a 24-mesh screen. The furnace charge
was 4 to 4J tons, and a strong fire was kept in order to quickly
ignite the ore. In about an hour, and before the sulphur com-
menced to burn, heavy arsenic fumes were given off, entirely
obscuring the interior of the furnace. After one and one-half or
two hours a sulphur flame could be observed to enter the flue,
which was a sign that the sulphurets were sufficiently ignited to
continue combustion without the aid of fire. The fire was then
allowed to go out and the fire and air-doors were kept wide open.
The temperature gradually increased by the combustion of the
sulphur until it reached a certain maximum, at which it remained
for several hours. After two or three hours the arsenic period was
over, the heavy fumes disappeared, the interior of the furnace
became clear, and the glow of the ore presented a beautiful pink
color. No chemical loss of silver took place during the arsenic
period. The oxidizing period continued for three to four hours
more, the ore remaining at about the same temperature through-
out, then fumes commenced to rise from the ore, and gradually
increased, but never became so dense as during the arsenic period,
though still dense enough to make the interior of the furnace
invisible. The chloridizing period had commenced, and the odor
130 HYDROMETALLURGY OF SILVER
of sulphurous acid and chlorine could be observed in the samples
taken during the first period; but later the odor of sulphurous
acid disappeared and only chlorine could be detected. In many
instances, however, no smell of chlorine could be noticed during
the whole time of roasting, as explained further on. Looking
through the open fireplace, which by this time had cooled down,
the whole interior seemed to be glowing, though the ore itself was
not visible on account of the zinc fumes, which were illuminated
by the glow of the red-hot ore and made a beautiful sight, and
at the same time afforded a good opportunity to observe the
temperature in the furnace. So soon as a decrease in temperature
was noticed, the fire was started again and kept up for three to
four hours, when the charge was considered finished. It took
altogether from twelve to fourteen hours to roast a charge.
During the oxidizing period the ore maintained an almost
level position in the cylinder and had a liquid-like appearance.
The ore particles on the surface, however, could be seen constantly
moving; on the side where the furnace moves up fresh ore came
to the surface as if emerging from a liquid, moved slowly across,
and sank as soon as it touched the down-moving side of the
cylinder. The ore increased considerably in volume during the
first part of the chloridizing period; but after it had reached
the maximum it commenced to shrink again, assuming a heavy
sandy condition, where before it was loose and woolly, and finally
occupied no more space in the furnace than the raw ore did.
A short time before the second fumes commenced to rise the
charge assumed a more inclined position, attaining nearly 45 deg.
There was a great and puzzling irregularity in the results.
Sometimes a number of successive charges gave satisfactory
results; then at once the chlorination dropped without any
apparent cause, the ore, the amount of salt, the temperature and
the treatment being unchanged. The only noticeable difference
was the amount of silver contained in the ore; and as a rule the
richer ore gave the better results, though it was evident the
quantity of silver as such could not influence the result materially.
Close investigation finally showed the true cause of the trouble
to be the more or less favorable proportion between the gangue
(carbonate of lime) and the sulphureted matters in the ore; and
this explanation corresponded with the fact noticed that richer
ores usually gave better results than poorer ores. No marked
CHLORIDIZING OF CALCAREOUS ORES
131
difference between the two could be noticed in an inspection of
the ore before it went into the battery.
The following tables of the results obtained with ores con-
taining less and ores containing more carbonate of lime illustrate
these differences:
ORE CONTAINING LESS CARBONATE OF LIME
(7 per cent, salt mixed in the battery.)
No. OF
CHARGE
VALUE OF RAW
ORE PER TON
VALUE OF ROASTED
ORE PER TON
VALUE OF LEACH
TAILINGS PER TON
CHLORINATION
Oz. Silver
Oz. Silver
Oz. Silver
Per Cent.
65
72.12
76.12
16.20
79.6
66
(a)
64.86
12.60
80.5
67
65.64
64.80
9.00
86.2
68
68.04
68.46
10.56
84.6
69
66.30
72.16
13.68
81.1
70
66.78
73.62
12.36
86.0
71
68.16
71.16
11.88
83.3
72
66.24
68.64
12.90
81.3
73
68.52
73.82
14.70
80.0
74
66.00
75.24
11.16
85.4
75
64.38
66.42
12.06
81.9
76
63.78
66.30
9.86
85.9
77
68.28
68.10
11.42
83.3
78
58.50
61.02
10.80
82.4
862.74
970.72
169.18
1161.5
(a) Sample of raw ore lost.
Average of raw ore 66.36 oz. per ton.
Average of roasted ore 69.33 oz. per ton.
Average of leach tailings 12.08 oz. per ton.
Average of chlorination 82.96 per cent.
The roasted ore contained 2.97 oz. more silver per ton than
the raw ore.
132
HYDROMETALLURGY OF SILVER
ORE CONTAINING MORE CARBONATE OF LIME
(7 per cent, salt mixed in the battery.)
No. OF
CHARGE
VALUE OF RAW
ORE PER TON
VALUE OF ROASTED
ORE PER TON
VALUE OF LEACH
TAILINGS PER TON
CHLORINATION
Oz. Silver
Oz. Silver
Oz. Silver
Per Cent.
79
61.06
60.36
15.90
73.7
80
62.46
64.20
19.14
71.2
81
60.18
64.08
16.14
74.9
82
56.70
56.52
11.76
79.2
83
55.08
55.44
12.72
77.1
84
57.12
53.88
20.80
61.4
85
60.34
58.98
14.52
75.4
86
59.64
57.60
20.04
65.3
87
58.50
56.10
18.18
67.6
88
49.92
52.86
15.18
71.3
89
55.38
54.90
12.84
76.7
90
55.80
52.56
12.90
76.2
91
57.60
51.36
9.12
81.3
92
57.00
55.80
13.20
76.4
93
55.02
55.74
15.00
73.1
94
58.32
56.64
14.64
74.2
95
57.00
57.36
7.62
86.8
977.12
964.38
249.70
1261.8
Average of raw ore 57.47 oz. per ton.
Average of roasted ore 56.72 oz. per ton.
Average of leach tailings 14.68 oz. per ton.
Average of chlorination 74.22 per cent.
The raw ore contained 0.75 oz. more silver per ton than the
roasted ore.
Comparing the average results, they are found to be decidedly
in favor of the higher sulphureted ore. To obtain further infor-
mation I concentrated some of the ore, and made mixtures of
certain proportions of concentrates and barren gangue, mostly
carbonate of lime, and roasted these different mixtures as well as
the concentrates in the muffle, treating all samples alike, using
7 per cent, of salt, and roasting each half an hour. The concen-
trates used were obtained from ore which did not represent the
average richness, assaying only 35.70 oz. per ton, and therefore
the different proportions I made were much poorer in silver than
I afterward found corresponding ones in the bulk of the run.
Only the concentrates were assayed, the values of the mixtures
being calculated.
These tests showed that the silver-bearing minerals of this
ore do not offer any difficulties to a good chloridizing roasting;
CHLORIDIZING OF CALCAREOUS ORES
133
on the contrary, they are easy to roast to a high percentage
without showing any tendency to ball. The difficulties the ore
offered were therefore caused by the gangue; and the great varia-
tion in the results was due to a greater or less favorable proportion
of sulphureted matters to lime.
No. OF
SAMPLE
CONCENTRATES
BARREN
GANGUE
VALUE OF
MIXTURE
PER TON
VALUE OF
LEACH TAILS
PER TON
CHLORINA-
TION
Per Cent.
Per Cent.
Oz. Silver
Oz. Silver
Per Cent.
1 (a)
100.0
96.0
2.91
97.0
2(6)
75.0
25.0
72.0
5.38
92.6
3(c)
62.5
37.5
60.0
4.72
92.2
4(d)
50.0
50.0
48.0
5.38
88.8
5(e)
25.0
75.0
24.0
5.47
77.2
(a) Fuming profusely during chloridizing period; strong and pure smell of
chlorine; a great deal of base-metal chlorides soluble in water was formed;
no tendency to form lumps shown. (6) During chloridizing heavy fumes;
distinct but moderately strong smell of chlorine; no base-metal chlorides
formed; no tendency to form lumps shown, (c) Much less fumes; very
little chlorine observable ; none in water-soluble base-metal chlorides ; no
tendency to form lumps, (d) Very little fumes; no chlorine; no base-metal
chlorides; showed tendency to form lumps, (e] No fumes except at a high
heat, and then but little; no chlorine ; none in water-soluble chlorides;
showed much tendency to form lumps.
Carbonate of lime in presence of heated sulphureted minerals
will change partly into calcium sulphate, which does not act on
sodium chloride, and partly into caustic lime, which decomposes
the metal sulphates and chlorides, and also, though less rapidly,
the silver chloride. If, however, carbonate of lime is greatly in
excess, only a very small amount, if any, of base-metal sulphates
is formed to decompose sodium chloride, the greater part being
changed directly into oxides. If no salt is present these sulphates
are, of course, quickly changed into oxides. For this reason no
chlorine can be detected if the salt is added after the ore has been
subjected to a partial oxidizing roasting.
Some of the more calcareous charges were followed closely,
but at no stage of the roasting could any soluble salts or chlorine
be detected if the salt was added after the ore had been oxidizing
for some time. If the salt was pulverized with the ore in the
battery traces of such salts were found, but never in large quanti-
ties.
As there were neither sulphates, which at a high temperature
134 HYDROMETALLURGY OF SILVER
decompose salt, nor chlorides, which at a high heat give off
chlorine, an increased temperature during the last part of the
roasting could not be of any benefit. In fact, it had even a very
bad effect. The caustic lime, which at a low temperature seemed
to be comparatively indifferent to silver chloride, decomposed it
energetically at a high temperature. The roasting of such ores
had therefore to be finished at as low a temperature as possible,
contrary to the usual practice. This fact must be taken into
consideration in constructing long reverberatory furnaces. The
arch of the first and second hearths nearest to the fire should be
very high. (Figs. 4, 5 and 6.)
Charge No. 86 of the more calcareous ore was roasted for
eight hours without fire; then, when the ore commenced to lose
heat, a second fire was applied for three hours. Before starting
the second fire a sample was taken through the entire length of
the furnace, and the chlorination was found to be 73.4 per cent.,
and after three hours' second fire it was only 65.3 per cent.
The moderately increased temperature at that stage of the
roasting reduced the chlorination 8 per cent. A number of such
observations were made and finally the mode of roasting was
changed, and the ore was allowed to roast entirely by the oxidation
of the sulphureted matters without the application of a second
fire.
Highly calcareous charges, containing only about 20 to 30
per cent. sulphurets,had a dead and sandy appearance, no chlorine
could be noticed, and it required a high heat to make the ore
fume. These fumes were light and were probably zinc oxide.
Not even at a very high heat could any chlorine be detected.
If, however, a very strong draft of air was allowed to pass through
the furnace from the beginning and during the whole time of
roasting, some chlorine was generated and could be detected.
The ore fumed moderately at a low temperature and 70 to 75
per cent, of the silver was rendered soluble in sodium hyposulphite.
This indicated that by a rapid supply of air more sulphuric and
less sulphurous acid was generated, and that some of the sodium
chloride was decomposed by the former.
The beneficial effect of a great surplus of air was still more
noticeable with highly sulphureted ore in the reverberatory furnace.
Differences as high as 25 to 30 per cent, in chlorination were ob-
served between two charges of the same ore, which were treated
CHLORIDIZING OF CALCAREOUS ORES 135
alike, except that one furnace had excessive draft while the
other had insufficient air. Silicious ores can be well chloridized
with a moderate supply of air, while calcareous ores need an excess.
This fact is of great importance, and due attention should be paid
to it in constructing furnaces.
An ore which by hand concentration showed 40 to 45 per cent,
sulphurets generated considerable chlorine; it fumed more freely
and at a lower temperature; and though no base-metal chlorides
were formed, 80 to 86 per cent, of the silver was chloridized when
temperature, time, and supply of air were properly regulated.
Ore showing by hand concentration 50 to 55 per cent, of sulphurets
permitted a chlorination of 90 per cent, and over.
In the Bruckner furnace the ore behaved very differently,
according as the salt was added in the battery or in the furnace,
and it was therefore necessary to examine each case separately.
ADDING THE SALT IN THE FURNACE
If the ore was allowed to oxidize without salt a certain per-
centage of the silver would be rendered soluble in sodium hypo-
sulphite, but not in water, owing to the arsenic contained in the
ore. This change takes place principally at the very beginning
of roasting and at a very low temperature. A sample taken
from charge No. 128 an hour after the combustion of the sul-
phurets had commenced, while the arsenic was still fuming
strongly and oxidization had but just commenced, the appear-
ance of the sample being still that of raw ore, yielded no Jess than
44.58 per cent, of the silver with sodium hyposulphite. Con-
tinuing the oxidation, the amount of soluble silver increased,
but not so rapidly as in the first hour, as was shown by the follow-
ing tests:
Charge No. 128, after 1 hour oxidizing roasting, yielded 44.58
per cent, of its silver with sodium hyposulphite.
Charge No. 128, after 1J hours oxidizing roasting, yielded
43.44 per cent, of its silver with sodium hyposulphite.
Charge No. 128, after 2J hours oxidizing roasting, yielded
45.32 per cent, of its silver with sodium hyposulphite.
Charge No. 128, after 3£ hours oxidizing roasting, yielded
57.66 per cent, of its silver with sodium hyposulphite.
No marked increase in soluble silver could be noticed by con-
tinuing the oxidation still further.
A certain percentage of silver can be extracted from some
136 HYDROMETALLURGY OF SILVER
ores with sodium hyposulphite without roasting, and in order to
determine whether this extractable form of silver was originally
contained in the ore, or was formed during even such a slight
oxidizing roasting as charge No. 128 showed, where in one hour
44.58 per cent, became extractable, a sample of raw ore was
lixiviated, but not even a fraction of 1 per cent, of silver could be
extracted. As this silver combination soluble in sodium hypo-
sulphite was principally formed in the first stage of oxidizing
roasting, during the time the arsenic was being expelled, it was
undoubtedly silver arsenate.
When the salt was added at any time during the oxidizing
period and well stirred in with a hoe, the percentage of soluble
silver either increased suddenly a few per cent, or decreased, but
in neither case could this percentage be maintained or further
increased. Decomposition commenced immediately and the per-
centage of extractable silver diminished rapidly, when, after two
to four hours, according to the amount of salt added, the maxi-
mum was reached.
In the following tables are given the detailed history of three
charges, roasted with 10, 7, and 4 per cent, of salt respectively.
The maximum of decomposition was reached in two, three, and
four hours:
CHLORIDIZING OF CALCAREOUS ORES
137
CHARGE NO. 116
(10 per cent, salt added during roasting.)
DESCRIPTION (a)
VALUE OF
ROASTED ORE
CONTAINING
SALT
PER TON
VALUE OF
LEACH TAIL-
INGS
PER TON
SOLUBLE
SILVER
Raw ore Oz. per ton, 62.52
Oz. Silver
Oz. Silver
Per Cent.
After 5 hours and 20 minutes roast-
ing and just before adding salt
Oz per ton 61 32
The same calculated to contain
10 per cent salt. .Oz.per ton, 55.19
Directly after adding salt and
stirred
5724
28.80
26 04
53.1
54 5
Twenty minutes after adding salt . . .
One hour after adding salt
Two hours after adding salt
62.04
61.98
63 24
31.92
39.48
46,68
48.6
36.3
26.2
Three hours after adding salt
Four hours after adding salt
Five hours after adding salt
62.10
64.32
64.20
42.90
41.64
40.02
31.0
35.3
37.7
(a) After 5 hours and 20 minutes oxidizing roasting 10 per cent, salt was
added and the charge stirred with hoes. The maximum of decomposition
was reached in 2 hours after the salt was added, at which time 51.92 per
cent, of the soluble silver was rendered insoluble. A decomposition of 10.82
per cent, took place in the first 20 minutes; in the next 40 minutes, 22.57
per cent.; in the next 60 minutes, 18.53 per cent.
138
HYDROMETALLURGY OF SILVER
CHARGE NO. 123
(7 per cent, salt added during roasting.)
DESCRIPTION (a)
VALUE OF
ROASTED ORE
CONTAINING
SALT
PER TON
VALUE OF
LEACH TAIL-
INGS
PER TON
SOLUBLE
SILVER
Raw ore Oz. per ton, 58.62
Oz. Silver
Oz. Silver
Per Cent.
After 4 hours roasting and just be-
fore adding salt . . Oz. per ton, 59.70
21 36
64 27
One hour after adding salt
59 84
27 00
*4 7s
Two hours after adding salt
Three hours after adding salt
56.04
55.92
35.40
41 16
34.69
26 40
Four hours after adding salt
54.60
3408
37 59
Five hours after adding salt
54.00
30.96
42 65
Six hours after adding salt . . .
57.06
27.12
52.48
Seven hours after adding salt. . .
50.52
23.10
54.28
Eight hours after adding salt
55.20
20.04
63.70
(a) The charge, subjected for 4 hours to oxidizing roasting before 7 per
cent, salt was added, yielded 64.27 per cent, of the silver soluble in sodium
hyposulphite. The maximum decomposition was reached in 3 hours after
the salt was added, at which time 58.92 per cent, of the soluble silver was
rendered insoluble. The decomposition during the first hour was 14.81 per
cent.; the second hour, 31.21 per cent.; the third hour, 12.90. During the
next 5 hours we find the amount of soluble silver gradually increasing until
at the end of the fifth hour nearly the same amount of silver was rendered
soluble as the ore contained before the salt was added. This would indicate
that by a continuation good results may finally be obtained, but the heat of
the charge is exhausted before that time, and a second fire causes decomposi-
tion again.
CHLORIDIZING OF CALCAREOUS ORES
139
CHARGE NO. 125
(4 per cent, salt added during roasting.)
VALUE OF
VALUE OF
ROASTED ORE
DESCRIPTION (a)
CONTAINING
LEACH TAIL-
SOLUBLE
SALT
INGS
SILVER
PER TON
PER TON
Oz. Silver
Oz. Silver
Per Cent.
Raw ore Oz per ton 59 70
After 3 hours roasting and just be-
fore adding salt . .Oz. per ton, 58.32
26.58
54.42
Just after adding salt and stirred . .
52.44
20.84
61.79
One hour after adding salt
56.16
23.94
57.38
Two hours after adding salt
56.10
29.40
47.60
Three hours after adding salt
54.90
42.24
23.07
Four hours after adding salt
58.56
46.80
20.09
Five hours after adding salt
56.50
39.66
29.25
(a) After 3 hours oxidizing roasting 54.42 per cent, of the silver in the ore
was soluble in sodium hyposulphite, then 4 per cent, salt was added and
mixed with the ore by hoes. A sample then taken through the entire
length of the furnace showed 61.79 per cent, soluble silver, or a gain of 7.39
per cent, in these few minutes. Decomposition soon set in, however, and
the maximum was reached in four hours after the salt was added, at which
time 67.48 per cent, of the soluble silver was rendered insoluble. The
decomposition during the first hour was 7.13 per cent; the second hour, 15.83
per cent.; the third hour, 39.70 per cent.; the fourth hour, 4.82 per cent.
Summing up the results of these experiments it was clear
that (1) at a very early stage of oxidizing roasting quite a high
percentage of the silver was converted into a combination (un-
doubtedly silver arsenate) which was soluble in sodium hypo-
sulphite, and which seemed to resist well the decomposing action
of the lime. (2) A part of this soluble silver was decomposed by
the action of the salt, the decomposition commencing almost
immediately after the addition of salt and continuing until a
maximum was reached, when a reaction took place by which
soluble silver was again formed — most likely silver chloride.
(3) A larger percentage of salt produced the decomposition
quicker, but not so thoroughly as a smaller percentage, viz. :
Charge No. 116, with 10 per cent, salt, reached the maximum
in 2 hours, with 51.92 per cent, of soluble silver decomposed.
Charge No. 123, with 7 per cent, salt, reached the maximum in
3 hours, with 64.27 per cent, of soluble silver decomposed.
Charge No. 125, with 4 per cent, salt, reached the maximum
in 4 hours, with 67.48 per cent, of soluble silver decomposed.
140 HYDROMETALLURGY OF SILVER
The singular fact was developed that when salt was added in
the reverberatory furnace during the oxidizing period no such
decomposition took place. Lixiviating tests made hourly and
half-hourly after the salt was added showed a gradually increas-
ing chlorination. This strange behavior of the ore in the Bruck-
ner seems to be a reaction of silver arsenate and sodium chloride,
but why such a reaction does not take place in the reverberatory
I cannot explain. No lumps or balls were formed in this mode
of roasting, but the peculiar reaction prevented a higher chlorina-
tion and rendered this method of roasting impracticable.
ADDING THE SALT IN THE BATTERY. — SELF-ROASTING
When the salt is added in the battery it becomes thoroughly
mixed with the ore, and both enter the furnace together. Though
there is also a considerable part of the silver converted into a
salt soluble in sodium hyposulphite at a very early stage of roast-
ing, due to the arsenic in the ore, yet no decomposition of the
arsenate of silver seems to take place. On the contrary, a gradual
increase of soluble silver is observed up to a very advanced stage
of roasting, when all the sulphureted minerals are converted into
oxides. If the roasting is continued beyond this point, or if
the temperature is raised before or after this point is reached,
then the caustic lime acts on the silver chloride and decomposes it.
When beginning the experiments I conducted the roasting
in the Bruckner the same as with certain highly sulphureted
silicious ores, viz.: (1) by raising a strong heat to start combus-
tion; (2) oxidizing without fire until the temperature commenced
to decrease; (3) chloridizing and finishing at a higher tempera-
ture with a second fire. In the course of the experiments, how-
ever, the employment of the second fire caused the decomposition
of a considerable part of the soluble silver already formed. A
charge was therefore roasted without using a second fire, leaving
the ore to complete the roasting in the heat created by the com-
bustion of the sulphurets, and much better results were attained.
For the sake of comparison I roasted and finished a number of
charges without the additional fire and then a number with the
second fire, using ore of the same grade and maintaining simi-
larity in all other conditions, and found that with self-roasting
the average chlorination was 5 per cent, higher, while the consump-
CHLORIDIZING OF CALCAREOUS ORES 141
tion of wood was reduced more than one-half. With a second
fire the average value of the roasted ore was 0.46 oz. silver less
than the value of the corresponding raw ore, while in self-roasting
the average value of the roasted ore was 1.91 oz. silver higher
than that of the raw ore, indicating smaller loss by volatilization.
I call this mode of roasting "self-roasting," because the ore,
once ignited, requires no further attention, and is left entirely
to itself until the heat has nearly died out. Self-roasting sim-
plifies the manipulation. It is only necessary to maintain fire
for about two hours, when the charge can be left to itself until
the time of discharging, which is about an hour before the red
heat dies out. Thus one man can attend to quite a number of
furnaces. Care has to be taken, however, to make large charges,
as small charges do not maintain the heat long enough, and they
"freeze" before the roasting is completed.
Having obtained by this mode of roasting an average of 82.9
per cent, chlorination, with occasional results of 85 and 86 per
cent, (see Table) with calcareous ore which offered so many dif-
ficulties to chloridizing roasting, it seems that it may be possible
to chloridize properly nearly all kinds of highly sulphureted ores
in this way; but, as large charges work better than small ones,
it is advisable to have the revolving cylinders large enough to
hold 5 or 6 tons of ore.
By roasting in this manner more subchlorides are formed,
and less volatile chlorides expelled; this, however, does not inter-
fere much with the subsequent lixiviation. The advantages
gained will more than overbalance the slight extra expense
caused by the increased consumption of sulphur, and the refining
of a somewhat baser precipitate, especially if the diminished loss
of silver by volatilization, which loss is principally caused by the
expulsion of the volatile base-metal chlorides, is taken into con-
sideration. If the ore contains copper, an increased formation
of cuprous chloride will even be beneficial for the subsequent
extraction by lixiviation.
The consumption of wood in self-roasting was found to be
only one cord for each 10.6 tons of ore, while in roasting with a
second fire only 4 to 4.5 tons could be roasted with a cord of
wood.
142 HYDROMETALLURGY OF SILVER
BALLING OF THE ORE
Another noticeable difference in the behavior of the ore, due
to the addition of the salt in the battery or in the furnace, was
the formation of balls or lumps in the first case, while in the latter
instance the ore remained loose without forming balls, and when
discharged ran on the cooling floor like water. When the salt
was added in the battery it assumed a more solid form, did not
spread over the cooling floor, and did not dust, but contained a
great many balls. These balls originated during the early part
of the chloridizing period and at first were of the size of a pin's
head. Gradually they assumed larger dimensions, and when
the charge was finished the majority of them were from the size
of a pea to that of a walnut. They were smooth, hard, and heavy,
consisting of concentric shells, and were formed even if a second
fire was not used. By reducing the speed of the furnace from
2£ revolutions per minute to one revolution in If minutes, and by
adopting self-roasting, which was equivalent to roasting at the
lowest possible heat, the number and size of these hard balls
were greatly reduced, but were not altogether prevented.
Roasting during the chloridizing period was tried with an
intermittent motion. The furnace was allowed to make one
revolution and was then stopped for fifteen minutes, when another
revolution was made, and so on until the charge was finished.
This reduced but did not prevent the formation of balls. Then
the furnace was stopped entirely during the chloridizing period,
intending to allow the completion of the roasting process with-
out any further movement, though it extended the time required
for roasting. This method would very likely have had the de-
sired effect had not another difficulty made it impracticable.
The surface, of the ore commenced to harden, forming a crust,
which increased in thickness, was spongy and porous, and would
not have interfered with the lixiviation, but threatened that by
the time the charge was finished the whole mass might have
hardened, which would have caused great difficulty in discharg-
ing, and would also have endangered the furnace through the
whole mass clinging to one side and then dropping suddenly as
the furnace revolved.
Though annoying, this formation of hard balls would not be
a serious obstacle, as by repeated tests it was found that they
CHLORIDIZING OF CALCAREOUS ORES 143
were well roasted; in fact, always a slight percentage better
chloridized than the fine stuff. By separating the coarse from
the fine, and crushing the former in a ball-mill or through rolls,
the ore would be well prepared for lixiviation. This extra hand-
ling does not signify much when the great difference in expense
between roasting in a Bruckner and roasting in a reverberatory
furnace is taken into consideration, to say nothing of other
advantages. Other lumps were formed toward the end of the
operation which were larger, but soft and porous, had a rough
surface, fell apart when brought in contact with water, and did
not interfere with lixiviation.
The sulphureted minerals of this ore had no tendency to cake
and form lumps. When the concentrates, free from gangue,
were roasted, the pulp remained perfectly loose and sandy, even
if only occasionally stirred during roasting; the formation of these
small hard balls must therefore have been caused by the gangue.
They were soft while hot and were readily crushed, but when
cold became hard and brittle, and when broken showed concentric
layers. They were not caused by excessive heat, because they
formed even when the heat was kept as low as the combustion
of the sulphureted minerals permitted; neither was their appear-
ance that of overheated ore. They were probably caused by the
formation of a double salt of calcium sulphate and sodium sulphate
(glauberite), Na2SO4 -f- CaSO4, a salt which fuses easily. This
would explain the singular fact that no lumps were formed if the
salt was added in the furnace after the charge had been oxidizing
for some time; in which case the salt could not be thoroughly
mixed with the ore, and therefore did not come in such close
contact with the sulphureted matters as the lime did, on which
sulphuric acid acts so much more energetically than on sodium
chloride that very little, if any, of the sodium chloride was con-
verted into sulphate. For the same reason free chlorine could
be detected only in exceptional cases if the salt was added in the
furnace. Neither was it possible to produce a chlorination of
the silver. The 40 to 60 per cent, soluble silver was arsenate of
silver.
In my muffle experiments with different mixtures of concen-
trates and gangue (limestone) the tendency to form lumps com-
menced with the proportion of 50 concentrates to 50 gangue,
and this tendency increased with the percentage of gangue; it did
144 HYDROMETALLURGY OF SILVER
not exist if the mixture contained more sulphurets than gangue,
and if the ore had been assorted at the mine to meet this require-
ment the formation of balls would doubtless have been entirely
avoided, the chlorination would have been much better, and the
Bruckner furnaces could have been used instead of the more
expensive reverberatories. It may be mentioned that the ore
from the mine had to pass successively through three large ore-
bins before reaching the mill. These were always kept full, in
order to have ore in reserve, and it took 10 to 14 days from the
time the ore was dumped into the bin at the mine before it reached
the battery; quick changes for experimental purposes were
therefore impossible.
ROASTING IN THE REVERBERATORY FURNACES
There were in all eight reverberatory furnaces in operation,
of the following dimensions:
Four furnaces, 50 ft. long by 10 ft. wide, containing 5 hearths each.
Two furnaces, 40 ft. long by 10 ft. wide, containing 4 hearths each.
Two furnaces, 30 ft. long by 10 ft. wide, containing 3 hearths each.
The arch at the highest place near the fire was 27 in. above
the hearth, further away only 20 in., and at the last hearth 18 in.
The sides were 10 in. high. The furnaces were built in pairs,
placed back to back, and each hearth had but one small working
door 8 x 12 in.
While tolerably good chlorination may be obtained with
silicious silver ore in furnaces of the above-described construction,
they surely were not of suitable design for the roasting of calca-
reous silver ores, as in the first place insufficient provision was
made for a free and well-located air inlet; the small working door
furnished air only for the ore near it, which fumed, while further
in and toward the back of the furnace it presented the appearance
of an inactive glowing mass of high temperature, and did not
emit any visible fumes. In the second place, the arch, especially
the one over the hearth nearest to the fire, was not high enough
to keep the temperature low enough for calcareous ores during
the chloridizing period. There was no remedy for these defects
except reconstruction. By keeping the fire-door continually wide
open much better results were obtained, but they were not
satisfactory until the furnaces were reconstructed.
CHLORIDIZING OF CALCAREOUS ORES 145
At Yedras I was astonished to find that the ore in the rever-
beratory furnaces on the finishing hearth was subjected to almost
a white heat, while care was taken to avoid cooling the furnace
by excluding the air as much as possible. The fire-door was
kept closed, and the very small working doors, of which there was
one on each hearth, were closed as soon as stirring was completed.
One of the charges roasted in this way consisted of one ton of
ore; four men were attending the furnace, which was 50 ft. long,
and 5 per cent, salt was added on the third hearth after the ore
had been nearly five hours in the furnace. The following is a
detailed record of same:
1. Raw ore without salt. Sample of ore taken from the
hopper before charging gave 67.08 oz. silver per ton.
2. Same charge. Sample taken from the first hearth after
the ore had been two hours in the furnace, and just before the
charge was moved to the second hearth, the hearth being dark,
gave off a pretty strong smell of sulphurous acid and contained
65.40 oz. silver per ton of ore.
3. Sample taken from the second hearth, after having been
there for two hours and twenty minutes, and just before the
charge was to be moved to the third hearth. Charge dark-red,
strong fume (arsenic period), with strong smell of sulphurous acid,
gave 66.06 oz. silver per ton of ore.
4. Sample taken from the third hearth just before the salt
was added. Charge red-hot, strong smell of sulphurous acid,
gave:
Leach tailings 31 44 oz | 53. 08 per cent, silver soluble in sodium hyposulphite.
5. Sample taken from third hearth after having roasted just
one hour with salt. Strong fumes and smell of sulphurous acid;
color, after cooling, light-brown; gave:
Leach tailings 29.04 oz! | 54'4 Percent" silver soluble in sodium hyposulphite.
6. Sample taken from fourth hearth after the charge had
been there one hour, and had roasted two hours with salt. The
ore woolly, and sample exposed to air fumed strongly; color, after
cooling, red-brown; temperature red; gave:
Leach tailings 28'.44 oz'. | 53 Per cent silver soluble in sodium hyposulphite.
146 HYDROMETALLURGY OF SILVER
7. Sample taken while charge was being moved to fifth hearth
and after roasting three hours and twenty minutes with salt.
Temperature light red; sample fumed strongly when exposed to
air; inside of the furnace clear, fumed only near the working door;
slight smell of chlorine and sulphurous acid; ore commenced to
assume a sandy consistency, and had the appearance of an over-
heated ore; color, after cooling, red-brown; gave:
Lelch tailings 27.12 oz'. } 61'8 Per cent" silver soluble in sodium hyposulphite.
8. Sample taken from fifth hearth after roasting five hours
with salt. No smell of sulphurous acid, very little of chlorine;
temperature very light red, almost white; ore sandy; color, after
cooling, greenish brown, gave:
Lelch tailings 23'.94 oz! | 65' 2 per cent> silver soluble in sodium hyposulphite.
9. Sample taken when the ore was discharged, after having
been eleven hours in the furnace and six hours roasted with salt.
Very little smell of chlorine, none of sulphurous acid; ore very
sandy, mixed with lumps; color, after cooling, greenish brown;
gave:
Leach tailings 2L06 oz! } 67'2 Per cent silver soluble in sodium hyP°sulphite.
The ore sustained during roasting a loss in weight by vola-
tilization of 13.8 per cent., and the loss of silver by volatilization
amounted to 16 per cent.
By comparing the different samples it was found that at the
time the salt was added there was 53.8 per cent, of the silver
soluble in sodium hyposulphite. This soluble silver was an
arsenate and was formed in the same manner as that obtained,
in the Bruckner furnace during the arsenic period. During the
next two hours of roasting with salt there was no change in
the soluble silver. This is unlike the behavior of the ore in the
Bruckner, in which during this period a marked decomposition
of the soluble silver took place. Although in the reverberatory
the salt was also added after the ores had been oxidizing in the
furnace for nearly five hours, and the arsenic period was over,
no soluble silver was decomposed by the salt. During the next
two hours it remained the same, and commenced to increase only
after the third hour with salt, and after the ore had reached the
CHLORIDIZING OF CALCAREOUS ORES 147
region of light-red heat on the fourth hearth. The increase,
however, amounted to only 8 per cent, and on the finishing
hearth, where the ore was exposed to an almost white heat for
two hours and forty minutes, a further increase of 5.4 per cent,
occurred — in all 13.4 per cent. — while the loss of silver by
volatilization, which took place principally during this period,
amounted to 16 per cent. Of the 67.2 per cent, of silver rendered
soluble in sodium hyposulphite, 53.8 per cent, was due to the
action of the arsenic, and only 13.4 per cent, to the action of the
salt. To gain 13.4 per cent, chlorination, 16 per cent, of silver
was sacrificed, a very thoughtless operation; and this was by no
means one of the worst results. With most of the charges 25 to
30 per cent., and even more, of the silver was lost, to gain a small
percentage in chlorination, as is shown further on.
The low chlorination was due to the insufficient supply of air,
which prevented the formation of sulphuric acid to satisfy the
lime and to act on the salt, while the great loss of silver was
caused by excessive heat and insufficient air. The different
behavior of the ore in the two furnaces as regards the decompo-
sition of the soluble silver after the salt was added is remarkable,
and is thus far unexplainable.
While experimenting with the Bruckner furnaces, the roasting
in the reverberatory furnaces was conducted in the manner just
described, and the experiments with these commenced after those
with the Bruckners were completed. One preliminary experiment,
however, was made, to demonstrate to the metallurgist in charge
the effect of the air; for he insisted that the Yedras ore could
only be chloridized at a very light-red heat, and he excluded the
air as far as possible, so as not to cool the furnace.
As mentioned above, the furnaces were not of proper construc-
tion, and the only means of getting more air into them was by
leaving the fire-door wide open. The beneficial effect of the air
was quite striking, though no change was made in the temperature,
the roasting proceeding at the same high heat as before. The
interesting record is given in the following tables:
148
HYDROMETALLURGY OF SILVER
ROASTING WITH 5 PER CENT. SALT AT A VERY HIGH HEAT
WITH CLOSED FIRE-DOOR
Reverberatory Furnace, No. 3
DATE
VALUE OF
RAW ORE PER
TON (a)
VALUE OF
ROASTED ORE
PER TON
SOLUBLE
SILVER
Loss OF SILVER
BY VOLATILIZA-
TION
January 13
14
Oz. Silver
63.03
62.04
Oz. Silver
56.52
50.04
Per Cent.
67.7
696
Per Cent.
17.8
25 5
15
16
17
18
19
20
21
63.78
63.78
64.08
61.20
60.60
61.56
60.00
49.44
52.80
60.00
49.08
45.00
49.50
48.48
57.1
49.8
43.0
43.2
63.2
68.8
68 1
27.2
25.4
13.9
26.3
31.7
26.1
25 7
22
58.20
45.48
76 8
278
Average
61.82
50.6*3
60.7
24.7
(a) Difference of value between raw and roasted ore, 11.9 oz. per ton.
ROASTING WITH 5 PER CENT. SALT AT A VERY HIGH HEAT
WITH OPEN FIRE-DOOR
Reverberatory Furnace, No. 3
VALUE OF
VALUE OF
Loss OF SILVER
DATE
RAW ORE PER
ROASTED ORE
BY VOLATILIZA-
TON. (a)
PER TON
SILVER
TION
Oz. Silver
Oz. Silver
Per Cent.
Per Cent.
January 23
57.66
49.56
85.3
20.6
24
55.50
47.04
79.7
21.6
25
54.36
50.40
79.6
14.3
26
63.30
55.20
83.9
18.0
27
63.84
59.22
77.7
13.9
28
64.86
61.08
73.5
12.6
29
65.52
57.00
65.5
19.2
30
66.78
59.40
77.2
17.4
31
66.00
58.80
74.5
17.3
February 1
62.10
55.68
79.6
16.7
2
57.30
53.40
75.7
13.7
3
54.54
48.96
76.2
16.6
4
55.20
52.80
77.5
11.2
5
57.60
51.96
75.8
12.0
6
60.66
55.14
80.5
15.6
7
61.68
53.70
72.1
19.2
8
64.50
65.64
68.9
5.5
9
55.80
55.56
75.5
7.6
10
57.96
48.48
80.7
22.3
11
62.46
56.52
76.8
16.0
12
67.02
59.94
69.5
10.7
Average
60.70
55.02
76.4
15.3
(a) Difference of value between raw and roasted ore, 5.68 oz. per ton.
CHLORIDIZING OF CALCAREOUS ORES 149
By comparing the two tables we find that from the very first
day the fire-door was kept open a marked improvement in the
results took place. The averages show this plainly:
Per Cent.
Average loss of silver — fire-door closed 24.7
Average loss of silver — fire-door open 15.3
Difference in favor of open fire-door 9.4
Per Cent.
Average of soluble silver — fire-door open 76.4
Average of soluble silver — fire-door closed 60.7
Difference in favor of open fire-door 15.7
Thus a reduction of 9.4 per cent, in the loss of silver and an
increase of 15.7 per cent, in soluble silver, or a total saving of
25.1 per cent, of the silver in the ore, was effected by simply
admitting more air into the furnace.
Preparatory to my experiments with the reverberatory fur-
naces I reconstructed some of them, to adapt them to the require-
ments of the ore. The arch was raised to 3 ft. with the exception
of the hearth furthest from the fire on which the ore was charged
and which was much lower. The working doors and the flue
were enlarged, and air-channels constructed to permit air to enter
along the fire-bridge and along the back wall of the furnace.
Roasting was then commenced at a much lower temperature,
and with a more liberal supply of air, much better results being
attained. The average chlorination of one month rose to 81.7
per cent, while the loss of silver was reduced to 1.7 per cent.
In the new furnaces the ore fumed over the whole hearth instead
of only near the working doors as before, and the temperature
was much more uniform in all parts of each hearth. To roast at
a low heat requires much greater attention and skill than at a
high heat, and it took some time before the men became accus-
tomed to it.
After numerous experiments to determine the proper tempera-
ture and draft, the following rules were adopted and will apply to
all calcareous sulphureted silver ores:
1. The fumes evolved must be kept in motion in all parts of
the furnace. If they stagnate around the ore, or if the furnace
assumes nearly a uniform heat throughout its entire length, it is
always a sign of insufficient draft, and if the draft is not increased
the result will invariably be a high loss of silver and a low chlori-
nation.
150 HYDROMETALLURGY OF SILVER
2. In no part of the furnace should the ore attain a light-red
heat. The fire should be regulated entirely according to the
temperature required at the finishing hearth nearest the fire,
where the temperature should be kept so that the ore has rather
a dark surface if not stirred, but a dark-red heat when the hoe
enters it. The ore, however, should fume, and the temperature
never go below the point required to evolve fumes. The presence
of fumes always indicates that the temperature is not too low,
while the dark surface of the ore shows that the temperature is
not too high.
An arch still higher than 3 ft. facilitates the maintenance of
the conditions prescribed in the second rule, and at the same
time permits a stronger fire for the more remote hearths without
injuring the charge on the finishing hearth. I rebuilt furnaces
Nos. 3 and 4 and made the arch of the finishing hearth 5 ft. high,
the sides 3 ft. 4 in., and the fire-bridge 2 ft. high; the next two
hearths were raised so that the arch was only 3 ft. high, the sides
16 in.; the next arch 27 in. high, sides 16 in.; and the charging-
hearth arch 24 in. high and sides 14 in. (Figs. 4, 5 and 6). These
gave the best and most uniform results of any of the furnaces;
the proper temperature on the finishing hearth being maintained
much easier, because the flame following the roof was so far
above the ore that it would have taken an excessive fire to over-
heat the charge. The chlorination obtained was very satisfactory,
and the loss of silver by volatilization was reduced to a minimum.
The average results of six weeks' working were: chlorination,
83.8 per cent; loss of silver, 0.8 per cent.
The consumption of wood in these furnaces was larger than
in the others: in the reverberatories with 3-ft. arch it amounted
to 0.098 cord per ton of ore roasted; and in the furnaces with a
5-ft. high arch 0.32 of a cord more per ton of ore was used; but
the better results, and the ease with which the temperature could
be controlled, overbalanced the extra consumption of wood.
When the roasting was properly conducted the percentage of
chlorination invariably depended, as in the Bruckner furnaces,
upon the proportion of sulphureted minerals and calcareous
gangue of which the ore was composed. The more sulphureted
matters the ore contained the better was the result. But little
difference was noticed in the formation of lumps whether the
salt was added in the battery or in the furnace; in both cases
CHLORIDIZING OF CALCAREOUS ORES 151
lumps were formed, but as a rule they were more porous and
softer than those of the Bruckner, and could easily be leached.
CONCLUSIONS
My experiments in roasting the calcareous and arsenical ore
of Yedras have shown:
1. That the main -difficulty in chloridizing the ore to a high
percentage was caused by the excess of lime in the ore. The ore
at Yedras was assorted, as in all silver mines, with the view to
obtaining a certain grade in silver, and no attention was paid to
the relative proportion of lime and sulphureted minerals, which
was of so great importance. In order to obtain regular chlori-
nations of over 90 per cent, it was necessary that the ore should
contain not less than 50 per cent, of sulphurets; but to accomplish
this it was necessary to break the ore smaller, so that more of
the barren limestone and calcspar could be thrown out, and
enable the sorters to pick out a poor grade of second-class ore.
Pieces containing much iron pyrites, even if they did not contain
silver, should be thrown with the first-class ore.
The accumulation of a second-class ore was not very desirable
at Yedras, because it could only be treated by concentration, and
the water supply was very limited, except during the rainy
season. However, I consider it more rational to accumulate a
second-class ore dump at the mine for future treatment than a
rich tailings pile at the mill, because it is very difficult to extract
the silver from tailings, while the chances are that after further
development the mine may produce more water; but even if it
should not, the accumulation of second-class ore would not be
more than could be concentrated during the rainy season. This
question is a very important one and due attention should be
paid to it. The outlay would be comparatively small compared
with the large amount of silver lost in the tailings.
By improvements in the mode of roasting I succeeded in
increasing the chlorination from 65 to 81 per cent, and above,
and in reducing the loss of silver from 25 to 1.7 per cent, and less,
thereby nearly doubling the production. If the ore had been
properly sorted the increase would have been at least 10 per cent,
more.
2. I proved that such ores could be roasted with a trifling
152 HYDROMETALLURGY OF SILVER
loss of silver by volatilization if the roasting was conducted as
described above, and that the enormous loss which the ore for-
merly sustained was caused by too high a temperature and an
insufficient supply of air.
3. I demonstrated that such ores, properly assorted, could be
successfully and very cheaply roasted in the Bruckner furnace
by the self-roasting process, if the salt is added to the ore in the
battery, provided suitable provisions be made to separate and
pulverize the hard but well-chloridized balls which form in these
furnaces.
PART II
EXTRACTION OF THE SILVER
XIII
LIXIYIATION WITH SODIUM HYPOSULPHITE
PERCY and Hauch were the first who proposed to convert the
silver in ores into chloride by roasting with salt and to extract
the silver with a solution of sodium hyposulphite, but Von Patera
was the first to apply this method in actual practice by treating
with it the rich and complex silver ores of Joachimsthal, Bohemia.
These ores were formerly treated by amalgamation and also by
the Augustin method, which both required chloridizing roasting;
but, on account of the great loss of silver which these ores sus-
tained during chloridizing, only the poorer grades were worked
by these methods. Patera, however, changed the mode of
roasting, inasmuch as he applied steam, whereby the loss of silver
by volatilization was much reduced, and made it possible that
even the richest varieties of the Joachimsthal ore could be suc-
cessfully chloridized. By using sodium hyposulphite as solvent
he introduced a process much cleaner and superior to either the
amalgamation or Augustin method. It was, however, executed
on a rather small scale; the roasting was done in charges of 500
to 600 lb., and the lixiviation was performed in tubs holding
about 200 lb. of roasted ore.
This process was introduced at Joachimsthal in 1858. Ten
years later, in the fall of 1868, I introduced it successfully at
La Dura, Sonora, Mexico, on a large scale, and in the following
years at Trinidad, Las Bronzas, San Marcial, in Sonora, Mexico;
then at Silver King, Arizona; Monitor, California; Cusihuiri-
achic, Chihuahua, Mexico, and at other places.
In course of time, as experience with different ores was acquired,
the process was much improved chemically, as well as with re-
gard to appliances.
The principle on which this process is based is the property
of silver chloride of being insoluble in water, while it dissolves
155
156 HYDROMETALLURGY OF SILVER
readily in a solution of sodium hyposulphite, even if the solution
is very dilute. To convert the silver, which in nature mostly
occurs as sulphide, into chloride, the sulphide ore is roasted with
salt (sodium chloride), as wre have seen in the first part of this
treatise. The solution dissolves only that part of the silver
which is converted into chloride, and the extraction depends,
therefore, entirely upon the quality of the roasting. The chlori-
dizing roasting is undoubtedly the most important part of the
process; and it cannot be too strongly impressed upon the minds
of operators to pay their principal attention to this operation,
and to make a thorough study of the behavior and characteristics
of their respective ores.
Before going into the details of the different operations it
will be more instructive to give a short description of the process.
In the first part of this treatise we have seen that, if a complex
ore be subjected to chloridizing roasting, not only the silver
sulphide but also the other metal sulphides are converted into
chlorides, subchlorides, sulphates and oxides. A number of these
salts are soluble in water and have to be removed from the roasted
ore by leaching with water before silver extraction, in order to
prevent them from entering the silver solution and afterward
the silver precipitate. This operation is termed "base-metal
leaching." The leaching with water, however, dissolves also
undecomposed salt, if such should be present, and the sodium
sulphate which was formed in roasting.
If by test it is ascertained that the soluble chlorides and sul-
phates have been removed from the ore, a diluted solution of
sodium hyposulphite is applied on top of the ore. This solution
by descending through the ore dissolves the silver chloride.
This operation is termed "silver leaching." The outflowing
solution is collected in special tanks (precipitation tanks). To
this silver solution a solution of calcium or sodium sulphide is
added, by which the silver and some other metals present are
precipitated as sulphide. By agitating the solution the precipi-
tate collects in dark-brown flakes, which settle quickly to the
bottom of the tank. When settled the clear solution is decanted
and returned to the process, and is used over and over again in-
definitely. The precipitate is filtered, washed, dried and charged
into a small roasting furnace to burn off the excess of sulphur
which it contains. This is done at a very low heat, and only as
' .
LIXIVIATION WITH SODIUM HYPOSULPHITE 157
long as the blue sulphur flame lasts. This free sulphur in the
precipitate is derived from the precipitant. The calcium and
sodium sulphides are both used as polysulphides. In precipitation
only one atom of sulphur combines with the metal, while the
balance drops down as such.
The calcined silver precipitate is finally melted on a lead-
bath in a cupeling furnace and refined, or it is shipped and
sold to smelting works.
After this synopsis of the process, I will proceed to describe
and discuss the different operations and appliances.
BASE-METAL LEACHING
The roasted ore is charged into wooden tanks, of which each
is provided with a filter bottom. The size of these tanks depends
on the intended working capacity of the works, the filtering
quality and the chemical nature of the roasted ore. Deep tanks
should never be used; in fact, the rule should be observed not to
make them deeper than 5 ft., and the diameter not to exceed
16 to 20 ft. Tanks with too large a diameter are difficult to dis-
charge. The reason why tanks should not be made deeper than
5 ft. will be given below.
Construction of the Leaching Tank. — Fig. 31 represents a
vertical 'and Fig. 32 a horizontal section of a leaching tank 16 ft.
in diameter and 5 ft. deep, made of 3-in. clear lumber. To use
lighter lumber than 3 in. is not an economical policy. The
lumber should be red wood or cedar, preferably the former,
because it is more indifferent to the action of chemicals than
any other soft wood. The bottom can be made of 12-in. plank,
but the staves ought not to be made over 6 in. wide, in order to
facilitate binding by the hoops. It is very important that the
tank should be tight in order not to lose by leakage valuable
silver solution, and therefore it should be made very carefully.
From the factory the parts are furnished in crated bundles and
the edges suffer considerably in transportation; besides, these
edges have usually only a rough saw-finish, and in putting the
tank together they ought to be planed by a skilled carpenter.
For this purpose several extra staves are always sent with the
shipment. The bottom planks should be kept in position by
wooden pins. Before erecting the tank the timber on which it
is going to set should be prepared, then the bottom planks placed
158
HYDROMETALLURGY OF SILVER
on top of it and pinned together, then the staves with their groove
slipped into the edge of the bottom and driven tight. The groove
in the staves should be f in. deep. The tank with its weight has
to rest on a circular bed of timber specially prepared, which must
be thick enough to leave the staves clear and free from weight.
Timbers 6x6 in. will answer; but they ought not to be further
apart than 12 in., as shown in Fig. 31. The best hoops are
those made in two or three sections of J or 1-in. round iron, and
!
I
/1'Qrove
ll* Dressed
^1* ^
>"£
1
Wro|ught
IroiJ Hoo
1& Pipe
FIG. 31. — LEACHING TANK, VERTICAL SECTION.
provided with cast-iron lugs for tightening. They resist chemical
action for a long time, while flat hoops are quickly destroyed.
It is well to paint them with two coats of asphaltum varnish
before placing them. One of them should be placed at a level
with the bottom, or only a very little below, so as not to interfere
with the outlet pipe of the tank. The hoops should be 12 in.
apart. The bed on which the tank rests should be made so that
the tank inclines toward the front 1 to 1£ in.
If a tank has been in operation for some time it will be found
that some material has accumulated under the filter bottom,
LIXIVIATION WITH SODIUM HYPOSULPHITE
159
which material has to be removed from time to time, so as to
permit a full flow of the filtrate from all parts toward the outlet.
To facilitate this cleaning the filter bottom should be made in
sections, small enough to be handled by two men. To the pieces
A, (Fig. 31), which are 4 in. high and 2i in. wide, 1 x 1 in.
laths are nailed 1 in. apart. Square notches cut into the pieces
FIG. 32. — LEACHING TANK, PLAN.
This shows filter bottom arranged in fourteen sections. Two sections are complete.
A permit the flow of the solution between the sections. Fig. 32
illustrates the arrangement of these sections, showing two of
them completed. All the lumber of which the filter bottoms
are made should be planed and painted with asphaltum varnish,
and the heads of the nails well driven into the laths, and covered
with white lead and asphaltum. The upper edges of the laths
160 HYDROMETALLURGY OF SILVER
should be beveled to prevent them from cutting the filter cloth.
For economical reasons it is well to order the laths together with
the tanks, and not attempt to make them at the mine.
About one inch above the filter bottom the tank is provided
with a groove. This groove is made by nailing first a wooden
strip 1 x 2 in. around the inner side of the tank, then by nailing
to this strip another 1x4 in. This gives a groove 1x2 in.
deep. It serves to tuck in the ends of the filter cloth. B, B,
Figs. 31 and 32, are the lead outlet pipes. It is not advisable to
use larger outlet pipes than \\ in. because they have to be con-
nected with a rubber hose about 5 ft. long, and if they are of
too large a diameter, flatten, and do not permit a flow in propor-
tion to their size, as the solution has no pressure. Over a \\-rn.
lead pipe a IJ-in. hose can be pushed, and if it is 5-ply will remain
round and discharge a full stream. If the solution flows out
with force it is a sign that the outlet is not large enough, and a
second or even a third one has to be provided. To check the
stream is a poor practice, because a free percolation is of great
advantage, as well with regard to the extraction as to the time
required for it.
Under but close to the top of the filter bottom is the J-in.
or f-in. lead pipe N ', Fig. 31. This pipe is connected with a
rubber hose which extends up to the rim of the tank. The hose
can be closed tight by a 'clamp-screw. This pipe serves to permit
the air from under the filter to escape. This is of importance if
the working of a tank has to be interrupted. If the solution
outlet is stopped, the air under the filter will be brought under
pressure by the solution, and, if there is no escape for it, it is apt
to break through the filter.
I have found the use of hose clamps in lixiviation works not
only very convenient but also economical. Valves are more or
less corroded by the solutions and are of short life, while the
clamp, doing the service of a valve, does not come in contact
with the solutions. These clamps are best when made of brass.
They are not in the market and have to be made to order. Figs.
33 and 34 give the construction and size of a clamp for a li to
2-in. hose. Smaller sizes for smaller hose have to be kept on
hand also.
The tanks ought not to be placed on a solid terrace, as we
often find done; it is much more rational to place them on
LIXIVIATION WITH SODIUM HYPOSULPHITE
161
trestles, because a leak in the tank is easily discovered, and,
besides, it gives room underneath to run cars for the removal of
the residues.
Solubility of Silver Chloride in Base-Metal Chlorides and Salt. —
The base-metal chlorides as well as sodium chloride have the
property of dissolving silver chloride, their dissolving energy
increasing with the temperature and concentration. It is there-
FIGS. 33 and 34. — BRASS CLAMPS FOR 1J-AND 2-IN. HOSE.
Full size.
fore not advisable to leach with hot water or to use too deep
leaching vats, even if the filtering quality of the ore permits it,
because the water by passing through a thick layer of ore becomes
highly charged with these salts and considerable silver chloride
will be dissolved. A very dilute chloride solution does not
dissolve silver chloride, and in order to prevent the base-metal
chlorides from dissolving silver, the washing of the ore should be
162 HYDROMETALLURGY OF SILVER
so conducted that a sufficiently dilute solution will be pro-
duced.
Prevention of Loss of Silver in Base-Metal Solutions. — In
tank lixiviation this can only be accomplished to a certain degree
(not perfect) by two methods which were introduced by myself:
(1) The water is allowed to enter the vat below the filter until it
gradually rises above the ore. Thus the most concentrated
portion of the solution will appear above the ore, and if then
diluted with a stream of water and the course of filtration reversed
a large part of the silver chloride will be precipitated and remain
in the ore, while the outflowing solution will be dilute, although
not sufficiently so as to be entirely free of silver. (2) A quicker
but not quite so effective a method is to fill the tank partly with
water and then to dump the ore into it, either dry or moist.
Enough water should be taken to rise about 4 in. above the
charge when complete. By this method the outflowing solution
will not be so dilute as by the first, but much more so than it
would be if the ore were charged in an empty tank and then leached
from above. However, the only method by which this problem
is actually solved is trough lixiviation, also introduced by me.
By this operation the ore can be brought at once in contact with
such quantities of water as to produce a solution so dilute that
not a trace of silver will be dissolved. More about this method
of lixiviation will be said further on.
The amount of silver dissolved during base-metal leaching
varies greatly, and depends on the nature of the ore, the thickness
of the layer, the amount of salt used in roasting, the mode of
treatment in the vat, and the temperature of the outflowing
solution. In some works, usually in large ones, the ore is charged
steaming hot from the cooling floor, or if for special reason it be
charged dry it will enter the tank even hotter, and as a matter
of course the first part of the outflowing solution will be very
hot and contain considerable silver. To illustrate how the
amount of silver dissolved by the base-metal solution varies
under different conditions, a few examples may be given:
San Francisco del Oro, Parral, Mexico. — Heavy zinc-lead ore
with iron pyrites, roasted with 4 per cent, salt ; the tanks partly
filled with water and the ore dumped dry and cool into it ; average
value of roasted ore 26.1 oz. per ton; charge, 8 tons; leaching
time, 8 hours; silver dissolved, 1 per cent.
LIXIVIATION WITH SODIUM HYPOSULPHITE 163
Sombrerete, Zacatecas, Mexico. — Heavy ore containing zinc
blende, galena, iron pyrites and some sulphureted copper minerals,
roasted with 6 per cent, salt; the tank partly filled with water
and the ore dumped dry and warm into it; average value of
roasted ore, 42.6 oz. silver per ton; charge, 52.5 tons; leaching
time, 12 hours; silver dissolved, 1.23 per cent.
Cusihuiriachic, Chihuahua, Mexico. — Not so heavy an ore,
containing galena, zinc blende, some iron and copper pyrites,
some silver-copper glance and also some ruby silver; roasted with
8 per cent, salt; the ore wetted on the cooling floor and charged
slightly warm into an empty tank and leached from above;
average value of roasted ore, 47.9 oz. silver per ton; charge,
8 tons; leaching time, 53 hours; silver dissolved, 2.5 per cent.
Determination of Amount of Silver Dissolved. — Mr. Russell
reports the amount of silver dissolved from the Cusihuiriachic ore
as 11.6 per cent, if the ore is charged cold and dry and leached
from above. Though Mr. Russell roasted with 10 per cent, of
salt I believe his figure to be much too high. His method of
ascertaining the amount of silver dissolved is not correct. He
determines on a sample in the laboratory the amount of salts
soluble in water contained by the roasted ore, and from this
figure he calculates the amount of silver which ought to be con-
tained in the ore after leaching with water, compares it with the
actual amount found in it by assay, and then estimates the
amount of silver dissolved. This is not correct, because a large
charge in the works cannot be so thoroughly washed as a small
sample in the laboratory, and therefore the calculated value of
the washed ore will be much too high, and consequently the
calculated amount of silver dissolved.
The correct way to ascertain the amount of silver dissolved is
to collect separately in vats the whole base-metal solution of one
charge and measure its volume. If it be then determined how
much silver is contained in 1000 c.c., it is an easy matter to
calculate the total amount of silver dissolved. To take the
sample out of the vats would not be correct, because the solution
being collected from different vats will not be uniform in concen-
tration and therefore will not be uniform as to tenor of silver.
Besides, a large portion of the silver will, be precipitated by the
gradual dilution of the first concentrated brine, which, to a
certain extent, will escape the sample, even if the contents of the
164 HYDROMETALLURGY OF SILVER
vat be agitated. However, there is no difficulty in obtaining a
correct sample. A small rubber tube, terminating with a glass
tube drawn to a fine point, is inserted in the outlet of the vat and
left there during the whole time of base-metal leaching. Thus a
very fine stream of the outflowing solution is obtained, and if
collected in a glass vessel of proper size, a representative sample
of about two or three gallons of the whole solution is obtained.
The volume of this sample is correctly measured and all the
heavy metal salts contained therein precipitated with calcium or
sodium sulphide. The precipitate is collected on a filter, dried,
weighed and assayed, and the amount of silver contained in it is
calculated. With this figure and the total volume of solution
collected in the vats the total amount of silver dissolved during
base-metal leaching may be estimated, and by knowing the
weight of the charge the amount dissolved per ton can be calcu-
lated and expressed in percentage. The figures of silver dissolved
under different conditions quoted above were obtained by this
method and therefore may be considered correct.
Effect of an Excess of Salt. — If the proper amount of salt be
used in roasting, so that all, or nearly all, be decomposed, and
the leaching be done by one of the improved methods, it may be
assumed that the amount of silver dissolved will never exceed
3 per cent. An excess of salt will cause a large percentage of
silver to be dissolved, reaching 60 to 70 per cent, in my own
experience. In this case the ore was of such a nature that it
required a large excess of salt in roasting to produce a satisfactory
chlorination of the silver. This very interesting occurrence will
be discussed below.
Precipitation of Silver from Base-Metal Solutions. — There
are several ways of precipitating the silver contained in the base-
metal solution, including precipitation by dilution with water,
to be described further on, and precipitation by one of the follow-
ing reagents: (1) Milk of lime. This produces a very voluminous
precipitate which is rather difficult to handle, and from which
the silver cannot be extracted unless there is a smelting furnace
connected with the leaching works. In such a case lime may
be used to advantage, especially since it is so cheap a reagent,
but care must be taken not to use the base-metal solution too
concentrated, in which case it is coagulated and much trouble
in further handling is caused. (2) Calcium or sodium sulphide.
LIXIVIATION WITH SODIUM HYPOSULPHITE 165
These precipitants are used in most of the lixiviating works, but
they are rather expensive, because they precipitate base metals
also. If sodium sulphide be added gradually to a tank charge
of base-metal solution, it will be observed that the precipitate
formed first is the richest in silver, and that it becomes poorer as
precipitation goes on. A complete precipitation, however, can be
produced only when all the base metals are thrown down. Since
the principal portion of the silver is precipitated in the beginning,
and only the smaller portion remains in solution with the main
bulk of the base-metal salts, the point will be reached in precipi-
tating when the cost of the precipitant will exceed the value of
the precipitated silver, and therefore a complete precipitation
is not advisable. In most lixiviation works only a partial pre-
cipitation is performed, after which the solution with the re-
mainder of the silver, which it did not pay to precipitate, is
allowed to run to waste.
Recovery of Silver from Waste Liquor by Precipitation with
Copper. — Although the waste liquor will be found very low
in silver, still if its large volume be taken in consideration it will
appear that the loss occurring in this way during a year is im-
portant. This loss, however, can be diminished greatly by pre-
cipitation with copper, which acts most energetically if finely
divided like cement copper, and if the solution be warm. To
effect a perfect desilverization it is necessary to treat the heated
solution in charges and to use mechanical appliances to produce
an intimate contact with the cement copper, but the base-metal
solution, after a partial precipitation with calcium or sodium sul-
phide, is not rich enough to warrant an expensive and complicated
treatment, and it will be found to be more profitable to employ
a more primitive method, even if not all of the silver is recovered.
The true economy in metallurgy, as in any other industry, is to
save the most at the least expense. As soon as its expense ex-
ceeds the value of the recovered metal, a method ceases to be
practicable, no matter how interesting it may be. That method
should be adopted which is best suited to local conditions.
The largest part of the dissolved silver can be saved by con-
veying the solution through a series of flat tanks in which cement
copper is so divided that it offers a large surface of contact to
the solution. The exhaust steam of the engine can be used to
increase the temperature of the stream, the steam being made
166 HYDROMETALLURGY OF SILVER
to enter the solution through pipes about 12 to 18 in. below the
surface in different tanks of the system. If the base-metal solu-
tion contains copper, the pipe projecting into the solution ought
to be of lead. The steam being condensed a vacuum is formed,
and very little if any back pressure to the engine will be noticed.
Wherever the climate allows it these tanks can be built in the
yard outside the works, without roof or shelter. If the ore con-
tains copper, sufficient cement copper will be formed by placing
scrap iron in the tanks, but if copper is wanting, the cement
copper has to be made from a dilute solution of blue vitriol.
Ores which are treated by lixiviation are seldom entirely free
from copper, and for this reason it is well to place some scrap
iron with the cement copper.
Base-Metal Leaching at Sombrerete. — At Sombrerete, Mexico,
I conducted the base-metal leaching in the following manner:
The ore is very permeable, and Stetefeldt and Russell, who built the
works, erected leaching vats 15 ft. 6 in. in diameter and 7 ft. 6 in.
deep, capable of receiving 55 to 58 tons of roasted ore. These
tanks were decidedly too deep. The ore filtered freely enough to
permit a deep charge, and in silver leaching this would not inter-
fere, but in base-metal leaching it caused much trouble. The
ore was not wetted on the cooling floor to produce additional
chlorination, but was charged dry. The water passing through 7
ft. of roasted ore became so saturated with salts that they crystal-
lized and blocked the filter, the space below the filter, and the
outlet pipe, which interfered greatly with the work and was very
annoying. I overcame this difficulty by filling the vats to the
depth of about 3 ft. with water and dumping the ore into it; fre-
quently the ore had to be charged while quite hot in order to
make room on the cooling floor, and consequently the outflowing
base-metal solution was warm and dissolved more silver chloride
than it would have done if more favorable conditions could have
been maintained. But notwithstanding this, the amount of
silver dissolved was much less than when the ore was leached
from above.
If a base-metal solution containing silver chloride be diluted
with water the silver chloride will precipitate as such. In the
Sombrerete works there were two base-metal precipitation vats
9 ft. 9 in. in diameter and 9 ft. deep; and in order to take advan-
tage of the above reaction I allowed the base-metal solution to
LIXIVIATION WITH SODIUM HYPOSULPHITE 167
run into both simultaneously, by which method I got the con-
centrated portion of the solution, which contained the most silver,
divided evenly into the two vats. This left considerable room
in them for diluting, and by allowing the less concentrated and,
finally, the very weak solution to run evenly into both vats, I
obtained a uniform solution dilute enough to cause the main
portion of the dissolved silver chloride to be precipitated. Silver
chloride thus precipitated being very finely divided requires a
long time to settle. For the purpose of effecting quick settling
and at the same time precipitating more silver, 5 to 10 gal. of
sodium sulphide solution were added to each vat and agitated.
In this way a part of the silver chloride was converted into sul-
phide, while the undecomposed part was readily collected by the
flocculent precipitate of copper, lead and other base metals, and
after agitation was interrupted the whole precipitate would
settle quickly.
After the precipitate has settled, the clear solution is decanted
by a stiff rubber hose, which enters the vat close to the bottom
and projects above the surface of the liquor. By lowering this a
few inches below the surface the liquor flows out, and if the hose
is lowered gradually as the liquor runs out it is always the clearest
part of the latter which is drawn off. This method of decanta-
tion is better than a series of tubes inserted at different levels in
the side of the vat, because during precipitation these tubes are
filled with precipitate, and in decanting the first solution from
each tube contains precipitate and consequently has to be con-
veyed to filters.
Outside the building were constructed a series of flat square
tanks 2J ft. deep, built of stone and mortar, well plastered and
coated with asphaltum varnish, which were provided with a
number of movable wooden double benches, or shelves, loaded
down with scrap iron. In this way the scrap iron was well dis-
tributed throughout the tanks and offered a large surface to the
solution. They were so arranged that the flowing solution
would move in its whole depth and had to take a zigzag course.
The ore of Sombrerete contains about 2 per cent, copper, and
the outflowing solution for a considerable time is colored green
by cupric chloride, and a good deal of copper is precipitated by
the scrap iron. The precipitated cement copper incrusting the
iron is loose and spongy and offers a very large surface for pre-
168 ' HYDROMETALLURGY OF SILVER
cipitating the silver. It ought not to be disturbed by stirring
the solution, since thereby the copper falls to the bottom and
does much less service. However, it is difficult to precipitate
all the silver, and it requires a rather large number of tanks.
Silver dissolved in a cupric chloride solution cannot be completely
precipitated until all the cupric chloride is decomposed; there-
fore as long as the solution leaving the last tank still gives a
reaction for copper it can be assumed to contain some silver also.
This indicates the necessity of erecting one or more additional
tanks. In Sombrerete these tanks were cleaned once a month,
and the cement copper obtained from them contained 500 to 600
oz. silver per ton, and 60 to 70 per cent, copper.
Cupric chloride in solution in contact with cement copper,
especially at an elevated temperature, is converted into cuprous
chloride, which settles as a heavy white crystalline precipitate.
Cuprous chloride again in contact with metallic iron is converted
into metallic copper, and the iron into ferrous chloride. It is
clear that in these tanks, in which the cement copper surrounds
the iron, an intimate contact between the iron and the cuprous
chloride is not possible, and that consequently the cement copper
will contain a large percentage of cuprous chloride. This is not a
desirable associate for the cement copper, because in the subse-
quent treatment it always causes a large loss of copper and silver,
and therefore the cuprous chloride ought to be decomposed.
This should be done shortly after the cement copper is taken
out of the tanks, because if left exposed to the air the cuprous
chloride will change into oxichloride. The quickest and most
rational way is to charge the cement copper in a revolving barrel
with an addition of water, salt and some light scrap iron. The
barrel is made of 3-in. staves and heads, and is bound with copper
hoops. It is constructed like an amalgamation barrel. The
inside is lined with hardwood lath about 1J in. thick, so that the
body of the barrel is protected from the wear and tear. Four
longitudinal ribs, projecting about 2J in., are inserted diametri-
cally in the lining, which produces a lively mixing of the charge.
The axles, or trunnions, which are made of brass, or better of
bronze, are provided with a strong flange by which they are
bolted to the heads of the barrel. Each is bored and provided
with a stuffing-box through which a copper pipe enters. One of
the pipes is turned downward close to where it enters, but must
LIXIVIATION WITH SODIUM HYPOSULPHITE 169
not project so far into the lower half of the barrel that the scrap
iron and the cement copper will strike it. Through this pipe
steam is introduced to heat the pulp. The pipe that passes
through the other trunnion is turned upward high enough to
just clear the ribs of the barrel. This pipe serves as outlet for
the gases and steam and leads outside the building. A manhole
on the side of the barrel, provided with a copper frame and lid,
serves for charging and discharging, while another in the head
gives entrance in case the lining is to be renewed or other re-
pairs are to be made.
When the barrel is charged with cement copper, water and
iron, salt is added, steam turned on, and the barrel is put in
revolution. A warm solution of salt (sodium chloride) dissolves
readily cuprous chloride, and from this solution the iron precipi-
tates the copper with great energy. Water should be used
moderately, just enough to produce a lively movement of the pulp,
say three to four times the volume of the cement copper, and to
produce a strong brine without using too much salt. The reaction
creates heat and the escape pipe has to be watched. When the
steam commences to come out freely the steam inlet ought to be
closed. The side of the barrel is provided with a plug-hole
through which the operator can take a sample by means of a
flask fastened to a copper wire. If the filtrate does not show a
reaction for copper the process is finished. This takes about 45
minutes. Below the barrel is placed a square flat filter tank,
above which is a large, strong iron screen with 2J-in. holes, the
frame of which rests on four car-wheels on rails, so that it can be
moved easily or withdrawn entirely from the tank. The con-
tents of the barrel are discharged on this screen, and the cement
copper washed into the filter tank by a stream of water, while
the scrap iron remains on the screen. Then some sulphuric acid
is added to remove the basic iron. This acid solution is allowed
to pass through the cement copper, after which the latter is well
washed. By this method I obtained cement copper containing
90 to 95 per cent, copper.
If the cement copper has to be prepared from blue vitriol it
may be desirable to use it for a longer time, in which case it will
become much richer in silver, but it has to be taken out occa-
sionally and treated with dilute sulphuric acid to remove the basic
salts. Thus purified the cement copper acts again with energy.
170
HYDROMETALLURGY OF SILVER
In order not to interfere with the regular work, the tanks
have to be so arranged that during cleaning half of them remain
in operation while the others are disconnected and cleaned. As
soon as half are cleaned, the scrap iron or cement copper is put
in place and the solution allowed to enter again while the other
tanks are cleaned.
Precipitating the Dissolved Silver Chloride by Dilution with
Water. — This method was first recommended and introduced by
me. All alkaline and metal chlorides have the property, when
concentrated, of dissolving silver chloride, and dropping it again
as such when diluted wTith water. The precipitation takes place
in proportion to the dilution, and if sufficient water be added all
the silver will be precipitated. This method of desilverizing the
base-metal solution is undoubtedly the most effective and cheapest,
all that is required being a sufficient supply of water and a few
more vats for base-metal precipitation than are usually found in
lixiviation works. I used this method first at the Silver King
mill in Arizona in 1880 to 1882 with very good results, then in
various other localities where the supply of water permitted it,
and also at the mill of the Hidalgo Mining Company, at Parral,
Mexico, in 1894.
Experience at Parral, Mexico. — The observations I made at
Parral were very interesting, and I shall give them in detail.
The ore which was treated consisted principally of galena, lead
carbonate, and blende, and was almost free from iron pyrites.
Neither galena nor blende produces in roasting with salt much
chlorine, especially if the blende belongs to that variety which
contains little or no iron sulphide. In this case the chlorination
of the silver depended principally on the chlorine produced by
the action of the quartz on the salt and by the direct action of
volatilized sodium chloride. Such roasting requires a large excess
of salt and a high heat. The roasting was conducted in a White-
Howell furnace with 9 to 10 per cent, of salt. Experiments were
made with the aim of reducing the proportion of salt, but as soon
as the amount fell below 9 per cent, the extraction suffered so
much that it was more rational to use a higher percentage of
salt. If the ore had contained sufficient iron pyrites 3 to 4 per
cent, of salt would have been sufficient, because lead-zinc ores in
presence of iron pyrites require less salt for a successful roasting
than any other class of complex ores; but since the iron pyrites
LIXIVIATION WITH SODIUM HYPOSULPHITE 171
was wanting and an excess of salt had to be used, and since zinc
blende and galena act but slightly on the salt, the roasted ore
contained a large amount of sodium chloride.
The lixiviation mill of the Hidalgo Mining Company is well
constructed and arranged, and reflects credit on J. T. Long, who
erected it. This mill has a very large cooling floor, on which the
roasted ore is allowed to cool dry for three days. Notwithstanding
this long time, the ore, when charged, is still hot enough to impart
to the outflowing base-metal solution for a rather long time' a
temperature of 140 to 200 deg. F. The excess of salt contained
in the roasted ore dissolved readily in the water of this temper-
ature, forming a highly concentrated brine in which 60 to 70
per cent, of the silver was dissolved, and therefore the principal
extraction was done during base-metal leaching. This, of course,
required a careful treatment of the base-metal solution.
Sodium sulphide was used as precipitant, but a large quantity
was required on account of the large amount of lead contained
in the solution. If a zinc blende ore containing iron pyrites is
roasted with the proper amount of salt and charged cool into the
vat the base-metal solution will contain but little lead, if any,
provided cold water is used; but if a concentrated hot solution
of sodium chloride is formed in leaching, the result is entirely
different, since this dissolves not only lead chloride, but also lead
sulphate, and the solution will contain large quantities of these
lead salts. If the solution be very concentrated, and allowed to
cool, large crystals will be formed, while the mother liquor re-
mains clear. If the solution be not very concentrated there will
be only a crystalline precipitate which turns the solution milky.
Neither the crystals nor the crystalline precipitate contain much
silver. The main part of the silver chloride, lead sulphate and
lead chloride remains in the solution, but if this be diluted with
water it turns milky, forming a heavy white precipitate; and if
sufficient water is used the white precipitate will contain all the
silver and lead dissolved in the solution before diluting.
For illustration I shall record my observations made on a
certain charge. The ore when charged was rather hot. The
leaching was done from above. The solution flowing out first
had a temperature of 200 deg. F. and was unusually concentrated.
Some solution was collected in a large beaker and left to cool.
A large amount of transparent crystals was formed. After pouring
172 HYDROMETALLURGY OF SILVER
off the clear mother liquor a quantity of water was added to the
latter, which caused a heavy white precipitate. This was sepa-
rated from the solution by filtration, and the three substances
were assayed for silver with the following results: (1) The crystals
contained 7.6 oz. silver per ton; (2) the white precipitate had
3386 oz. silver per ton; (3) the filtrate had none. Of this charge
of ore, which contained 19.6 oz. silver per ton, 62.6 per cent, of
the silver was extracted during base-metal leaching.
Working by this method I found it more convenient to do the
precipitation in the trough leading from the lixiviating vats to
precipitation vats, by allowing a stream of water to enter the
trough. In flowing some distance the solution and water become
thoroughly mixed, and the solution is perfectly desilverized before
it enters the vats. The stream carries along the white precipitate,
which settles in the precipitation vats. To ascertain whether the
solution receives the proper quantity of water or not, a sample
of the diluted solution just before it enters the vat is taken and
filtered. Of the filtrate 50 c.c. are poured into a beaker and
150 c.c. of cool, clean water are added. If it remains clear after a
few minutes it shows that no more water is needed; if it turns
milky the stream of water has to be increased. Thus by decreas-
ing or increasing the stream the proper dilution can be main-
tained. Of course the base-metal solution can also be diluted in
the precipitation vats, but by this method, if the two streams
simultaneously enter the vat, it is difficult to ascertain the proper
proportion, and if part of the vat is filled with solution and then
water is added there will be a considerable loss of time. Where
mechanical devices are used for agitation it is advisable to have
the solution agitated while the vat is filling; where agitation is
performed by hand, it ought to be done from time to time, to
make the precipitate settle more quickly. The tanks are filled
to about 12 to 18 in. below the rim, to leave room for more water
in case the proper proportion was not maintained in the trough
and a correction is necessary. When this is done a short time is
allowed for the heavy part of the precipitate to settle, then one
quart of sodium sulphide is added and the solution agitated
again. This is done to convert the very fine particles of the
precipitate, which otherwise would remain suspended for a long
time, into sulphide, in which state they assume a flaky condition
and settle much more quickly.
LIXIVIATION WITH SODIUM HYPOSULPHITE 173
The lixiviation tanks were charged with 30 tons of roasted
ore, and the base-metal leaching required about twenty-four hours.
If the precipitation was not done by water but by sodium sulphide,
the base-metal solution filled 7J- tanks of 3500 gal. each, and
53 gal. of the precipitant were used. The solution running to
waste still contained 0.19 oz. silver per 1000 gal., or 0.15 oz. per
ton of ore leached. If the precipitation was carried so far that
all the base-metals were precipitated, the solution running to
waste did not contain any silver, but then the cost of the precipi-
tant exceeded the value of the silver saved. If the precipitation
was effected by dilution with water in the trough, 22 precipitation
charges of 3500 gal. each were filled. The consumption of sodium
sulphide amounted to 2.75 gal. only, and the clear solution
running to waste did not contain any silver at all. This result
shows that it requires on an average about two parts of water to
one part of solution. Of course, in the beginning, when the out-
flowing solution is very concentrated, considerably more water
has to be added than in the above proportion, but the volume of
water to be added decreases as leaching progresses, and toward
the end it is less than the volume of solution.
On account of the large quantity of lead which the base-metal
solution carried, and which is precipitated by water as well as by
sodium sulphide, there was not much difference in the silver
contents of the precipitate whether obtained by water or by
sodium sulphide. In this particular case, however, a great differ-
ence in the silver contents of the precipitate was observed if the
water for leaching was first applied below the filter or from the
top. In the former instance it contained 1200 oz. silver per ton;
in the latter only 800 oz.
The precipitation with water is undoubtedly the cheapest and
most effective mode of treating the base-metal solution, and it
ought to be used in all works where there is an ample supply.
The Use of Cupric Chloride During Base-metal Leaching. —
The chlorination of the silver, and correspondingly the extraction,
can be increased greatly in ores which do not contain any or only
a little copper, by use of a solution of cupric chloride while base-
metal leaching is in operation. The cupric chloride can be pre-
pared in the works by boiling one part of blue vitriol with two
parts of salt in a wooden vat by direct application of steam.
The cupric chloride is made in large quantities and used as re-
174 HYDROMETALLURGY OF SILVER
quired. It takes from 3 to 4 Ib. of blue vitriol and 6 to 8 Ib. of
salt per ton of ore. To avoid the formation of too concentrated
a solution the cupric chloride is not added until the leaching has
proceeded one or two hours. It is better not to add the whole
of it at once, but to divide it so that one-fourth is added to about
8 in. of water above the ore; and when this charge is disappearing
below the surface another charge of 8 in. water and the second
fourth of the cupric chloride is made, and so on until all is used.
If the outflowing solution is green, it ought to be collected in a
separate vat, and can be used over again either for the same
charge or for the next one following, according to circumstances.
It is best to lift this solution by means of a lead steam siphon or
injector. If an ore contains so much copper that the base-metal
solution is colored green, this reaction, which produces an addi-
tional chlorination of the silver, takes place in every charge in
the regular operations of the process, and this is one of the reasons
why cupriferous ores are the easiest to treat and permit the
closest extraction. However, even with such ores, better results
are obtained if the green base-metal solution is collected separately
and allowed to pass through the charge a second time. Any
defect in the roasting will be much lessened, and uniformly good
results will be obtained. The more inferior the roasting the higher
will be the percentage of gain in chlorination. At Parral, where
I treated the San Francisco del Oro ore containing 25.5 per cent,
zinc, 11.56 percent, lead and 1.02 per cent, copper, on imperfectly
roasted charges by the use of cupric chloride 32 to 38 per cent,
in chlorination were gained. At the Silver King, Arizona, where
I treated an ore containing a large percentage of copper, I in-
creased the extraction a little by using part of the green solution
over again. With ore from the Veta Grande of Parral, which
contained mostly zinc blende, galena, and lead carbonate, and
no copper, I experimented on a large scale with cupric chloride,
and obtained much better and more uniform results.
SILVER LEACHING
After the ore is freed from base-metal salts by leaching with
water, a weak solution of sodium hyposulphite is introduced on
top of the ore. It is of great importance to work with a weak
solution. If the solution is strong, 1, 2, or 3 per cent., and an
attempt is made to maintain this as standard strength, it can be
.
LIXIVIATION WITH SODIUM HYPOSULPHITE 175
done only by frequent additions of sodium hyposulphite; else the
volume of the stock solution will decrease rapidly and the prepa-
ration of a new stock solution will soon be required, because the
ore absorbs a large volume of solution, and when, after the silver
extraction is concluded, this solution is displaced by water, it is
impossible to regain the same volume of the original strength.
It will be found that only a comparatively small portion of the
outflowing solution will have the standard strength, and if the
displacement of the solution is done so as to maintain the same
volume of stock solution, much water will enter the latter, while
considerable hyposulphite salt will remain in the residues. Such
a dilution of the stock solution takes place also in the beginning
of the silver leaching. The " hypo " solution passing through the
washed ore has to displace the water absorbed by the ore, and
this cannot be accomplished without the first portion of the
solution becoming mixed with the water; and since even a very
dilute solution of sodium hyposulphite dissolves silver chloride,
the outflowing solution has to be conveyed to the silver precipi-
tation vats a long time before it reaches its standard strength.
Therefore it will clearly be seen that the standard strength
can only be maintained by a continual addition of sodium hypo-
sulphite to the stock solution. A continuous supply is furnished
by the precipitant, whether that be sodium or calcium sulphide,
because in preparing it some sodium or calcium hyposulphite is
formed. As much as 5 to 7 per cent, of this salt may be found
in the freshly prepared precipitant, without the latter being
exposed to the action of the air for any considerable length of
time. The constant supply from this source, however, does not
suffice to replace the loss of hyposulphite salt which the stock
solution suffers if the latter has to be kept at a strength of 1 or
2 per cent., but it does suffice, and in fact in most cases exceeds
the loss, if a weak solution is used. The consequence is that in
the first case sodium hyposulphite has to be bought and added,
which increases the treatment expense, while in the second case
not a pound of this salt has to be added, even during years of
operation, and, on the contrary, water has often to be added to
reduce the strength to the standard.
The following table is an interesting daily record of the sodium
hyposulphite contained in the stock solution with which I worked
the San Francisco del Oro ore. With the exception of June 16,
176
HYDROMETALLURGY OF SILVER
on which day, for a certain experimental purpose, 175 Ib. of
sodium hyposulphite were added, there was not a pound of this
salt added to the stock solution:
2
2
2
2
Kz
K z
Kz
Kz
r. O
fe O
"•9
OH
0 H
0 H
OH
]
)ATE
H 2
DATE
. D
]
)ATE
H J
]
3ATE
H§
we/)
il
Z O
WC/3
If
(6 **
OS ""
C6 "
M "*
OH
w
0<
£
£
AF
>ril 13
0.34
May 11
0.54
Ju
ne 7
0.39
Ju
ly 4
0.76
14
0.52
" 12
0.56
8
0.39
5
0.70
15
0.50
" 13
0.56
9
0.39
6
0.69
16
0.55
" 14
0.56
10
0.40
7
0.73
17
0.60
" 15
0.52
11
0.40
8
0.78
18
0.58
" 16
0.49
12
0.38
9
0.78
19
0.56
" 17
0.53
13
0.38
10
0.79'
20
0.61
" 18
0.53
14
0.38
11
0.81
21
0.60
" 19
0.52
15
0.38
12
0.63
22
0.58
" 20
0.57
16
0.50
13
0.61
23
0.58
' 21
0.57
17
0.50
14
0.58
24
0.58
' 22
0.59
18
0.49
15
0.55
25
0.57
' 23
0.58
19
0.48
16
0.56
26
0.56
' 24
0.54
20
0.53
17
0.59
28
0.53
' 25
0.56
21
0.55
18
0.60
29
0.54
' 26
0.58
22
0.55
19
0.62
30
0.51
' 27
0.60
23
0.56
20
0.61
M
y 1
0.51
' 28
0.60
24
0.58
21
0.63
2
0.49
' 29
0.63
25
0.65
22
0.64
3
0.51
' 30
0.63
26
0-67
23
0.66
4
0.49
' 31
0.59
27
0.70
24
0.66
5
0.50
June 1
0.49
28
0.75
25
0.65
6
0.46
" 2
0.53
29
0.75
26
0.50
7
0.47
" 3
0.53
30
0.73
27
0.63
8
0.47
" 4
0.54
Ju
y i
0.80
28
(?)
9
0.50
" 5
0.48
2
0.81
29
0.59
10
0.49
" 6
0.43
3
0.79
It will be noticed that the solution had a tendency to increase
in strength. In order to keep it as much as possible at the stand-
ard strength of 0.50 per cent, water had to be added frequently,
which accounts for most of the sudden drops. A sample of the
stock solution taken after precipitation contained: hyposulphite
salts, 0.52 per cent.; lime, 0.165 per cent.; sulphuric acid, 0.140
per cent.; chlorine, 0.098 per cent. The solution evaporated and
heated in the muffle gave 0.775 per cent, solids. A month after,
it gave 0.785 per cent, solid residues. This shows how remark-
ably clean the solution remained. The 0.14 per cent, of sul-
LIXIVIATION WITH SODIUM HYPOSULPHITE 177
phuric acid corresponds exactly with a saturated solution of
gypsum (1 part in 400 parts of water). The 0.098 per cent,
chlorine is equal to 0.16 per cent, sodium chloride.
Best Strength of Solution. — In my experience I have found
the best strength of the solution to be 0.25 to 0.50 per cent.
Such a solution offers not only the above-mentioned economical
advantage, but it produces also a precipitate much cleaner and
richer in silver. Lead sulphate and cuprous chloride dissolve
much more easily in a strong than in a weak solution, and there-
fore with a strong solution a very low-grade precipitate will result
and a much larger amount of the precipitant will be required.
It is a wrong supposition that a strong solution shortens the
leaching time and extracts the silver better. The strong solu-
tion, dissolving readily lead sulphate and cuprous chloride, be-
comes heavily charged with these salts and thereby loses much
of its dissolving power for silver chloride, and therefore neither
the leaching time will be shortened nor will the extraction be
better.
Method of Leaching. — When base-metal leaching is finished and
the sodium hyposulphite solution is to be applied, it is well to do
this before all the water has disappeared from the surface of the
ore. The solution, th.en, in its course downward, follows closely the
water and replaces the latter perfectly in all parts of the charge,
and has less opportunity to become mixed with the water absorbed
by the ore than if the water is allowed to drain from the ore be-
fore the solution is applied, and consequently the reaction, when
the silver appears at the outlet, will be more precise. There is
another advantage connected with it, inasmuch that if the charge
be allowed to drain, the space before occupied by water will be
filled by air, and when the hyposulphite solution is run on, it is
very difficult, in fact almost impossible, to drive this air out again.
The layer next to the surface will give up its air, which can be
observed by the bubbles at the beginning, but further down in
the charge the friction becomes too great, and the air, not being
able to escape, will compress to let the solution pass, and by
doing so will reduce the speed of filtration and prevent a quick
and intimate contact between the ore particles and solution,
thus increasing the time of leaching and decreasing the percentage
of extraction. Therefore attention should be paid that during
no time of lixiviation the liquid be allowed to sink below the
178 HYDROMETALLURGY OF SILVER
surface of the ore. In a fresh charge the ore is loose, and when
water is applied on top it will, in descending, easily force down-
ward the air, which will escape through the outlet.
Testing the Solution for Silver. — It requires close watching
to ascertain the time when the silver appears at the outlet. If
the liquid is tested with sodium or calcium sulphide the reaction
is very uncertain, because the base-metal leaching is seldom
carried to such an extent that no light clouds are formed by an
addition of these reagents, and therefore the reaction cannot
show the first traces of silver in the outflowing stream. Even a
very dilute solution of sodium hyposulphite dissolves silver
chloride, and we can safely assume that, as soon as sodium hypo-
sulphite can be detected at the outlet, the stream contains silver;
and since furthermore the liquid naturally will contain more
hyposulphite salt than silver, it is safer to adopt a method by
which this salt can be detected in the outflowing stream. I use
the following test, which is reliable and convenient:
A small strip of starch paper is dipped into iodine solution
and then held in the stream. If the blue color disappears it is
a sign that the liquid contains a hyposulphite salt and conse-
quently silver, and the stream has to be turned at once into the
trough leading to the precipitation tanks. The base metals have
to be leached with cold water to make this test applicable, because
hot water also discolors the blue paper. I advise all who prac-
tise the lixiviation process to use this test and see that when
base-metal leaching is changed to silver leaching the outflowing
stream is very closely watched. By doing so they will find
that shortages in silver for which they could not account before
will be avoided.
Effect of Lead and Copper. — If the ore contains lead and
copper we shall find both metals in the hyposulphite solution,
because lead sulphate and cuprous chloride which are present
in the roasted ore are not soluble in water but dissolve in sodium
hyposulphite. Both these metals are precipitated together with
the silver, and we find them in large quantities in the precipitate,
reducing materially the fineness of the latter. To remove the
lead from the solution Mr. Russell precipitates it as carbonate by
adding sodium carbonate to the solution previous to precipitation
with sodium sulphide. By doing so he obtains a sulphide pre-
cipitate free from lead and lead carbonate as a by-product, but
LIXIVIATION WITH SODIUM HYPOSULPHITE 179
this complicates the manipulations without offering much prac-
tical advantage. Only where the sulphides are refined by melt-
ing them with iron and borax in crucibles is the use of this method
justified, because if the precipitate is free from lead a silver
bullion over 900 fine will result. But this treatment is too ex-
pensive and inconvenient, and is not used except in small works,
and then only in exceptional cases. In large works the refining
is done on a lead bath in the cupeling furnace, and lead in the
precipitate is then of great advantage.
For the sake of information the Russell method was tried in a
modified way. Instead of precipitating the lead in the solution
after it had left the lixiviation vat, it was precipitated inside the
ore charge by adding sodium carbonate to the solution before
leaching. The outflowing solution was entirely free from lead.
The precipitated lead carbonate remained in the ore. No in-
jurious effect on the extraction was noticed. The residues con-
tained about as much silver as when leached without an addition
of sodium carbonate, but the precipitate was much richer in
silver.
Calcium Sulphide as Precipitant, and Action of Calcium Hypo-
sulphite. — If calcium sulphide is used as precipitant the solution
will contain after precipitation calcium hyposulphite, and it
used to be generally assumed that within a short time the original
sodium hyposulphite solution was replaced by calcium hyposul-
phite. I discovered and demonstrated that this is not the case;
in fact, that even if the original lixiviating solution were calcium
hyposulphite, that compound could not exist in the process for
any length of time, but would be converted into sodium hyposul-
phite. We read in standard works that Patera leached with
sodium hyposulphite and precipitated with sodium sulphide,
while Kiss used calcium hyposulphite as solvent and calcium
sulphide as precipitant. It is mentioned that calcium hyposul-
phite possesses a greater dissolving energy for gold than the cor-
responding sodium salt, and that for this reason the Kiss process
is more suitable for gold-bearing silver ores. However, Kiss was
not leaching, as he thought, with calcium hyposulphite, but with
sodium hyposulphite, and the better extraction of gold was
most likely due to cooling the roasted ore slowly.
In roasting sulphureted ore with salt, sodium sulphate is
formed in large quantities. This salt is not easily removed from
180 HYDROMETALLURGY OF SILVER
the ore by leaching with water. If leaching with water be pro-
longed it will be found that the outflowing liquid will react for
sodium sulphate long after all the heavy metallic salts are re-
moved. Therefore, when the hyposulphite solution for the ex-
traction of silver is applied the ore still contains sodium sulphate
in considerable quantity. Calcium hyposulphite reacts with
sodium sulphate, forming sodium hyposulphite and insoluble
calcium sulphate. If, therefore, calcium hyposulphite is brought
into the stock solution by the precipitant, or when calcium hypo-
sulphite is used as solvent, it will be converted into sodium hypo-
sulphite by the regular operation of the process.
Time Required for Lixiviation. — The time of lixiviation varies
according to the nature of the ore, its permeability, and the size
of the charge. I observed that if the main part of the silver in
the ore is contained in galena, the silver extraction will be slow,
even if the ore filters well, while if all or the principal part of the
silver is contained in copper, zinc, arsenical or antimonial minerals
the extraction is quick. At the Silver King mine, Arizona, a
charge of 8 tons required only nine hours' silver leaching, though
it contained on an average 161.4 oz. silver per ton; while at Cusi-
huiriachic, Mexico, the time required for a charge of 8 tons was
fifty-three hours, the ore containing but 47 oz. silver per ton.
The filtering capacity in both cases was very nearly the same.
At the Silver King the silver was mostly contained in gray copper
ore, antimonial fahlerz and silver-copper glance, while at Cusi-
huiriachic it was principally contained in galena. I have made
the same observation with ores of many other localities.
A free filtration is very important for a quick and thorough
extraction. The ore particles should be brought rapidly in con-
tact with fresh solution, which cannot be done if the filtration is
slow. After the solution, descending through the ore, has dis-
solved a certain quantity of the salts present it loses much of its
dissolving energy, and therefore the small stream of a bad filter-
ing ore will not contain much more silver per liter than the large
stream of a quick-filtering ore. The same observation can be
made if the outflowing stream of a quick-filtering ore be checked
and reduced to a small stream. Likewise, if lixiviation is in-
terrupted and the solution allowed to remain in contact with the
ore for some time, say over-night, the solution will not contain
much more silver after leaching is resumed than the stream did
LIXIVIATION WITH SODIUM HYPOSULPHITE 181
before the interruption. Ores which after roasting run on the
cooling floor like water always filter badly and are not suitable
for tank lixiviation. Mixing such ores with sand or chopped
straw does not improve their permeability. By mixing in a
small percentage of galena before roasting, however, the ore
loses somewhat of its dusty condition and permits a little better
percolation. The only rational way, however, of treating such
ores is by trough lixiviation.
Regeneration of a Foul Hypo Solution. — The same sodium
hyposulphite solution was used at the Silver King works, Arizona,
for over a year and a half without requiring any addition what-
soever, of sodium hyposulphite, and acted still as energetically as
it did the first day. The ore permitted a very quick extraction
of the silver; the silver leaching did not require more than nine
to ten hours. Suddenly a change took place. The solution,
which formerly gained in strength and had to be diluted from
time to time, became weaker; the time required for silver leaching
increased to fifty, sixty, and seventy hours. The main portion
of the silver was extracted in twenty-four hours, but the balance
took a very long time. The value of the precipitate dropped from
$12 per pound to $6, and the leaching tanks, which always were
ahead of the furnaces, could not keep up with them. An addition
of sodium hyposulphite salt to the stock solution improved it
somewhat, but only for a short time. Seeking for the cause of
this strange change, I examined closely the present ore and
compared it with that which was worked before, as the cause
could be found only there, because the roasting as well as the
leaching was done in exactly the same way as during the previous
year and a half. I found that the present ore, though its general
appearance was the same as before, did not carry any copper.
This occurred when the works at the mine reached the 700-ft.
level. Not being able to find any other difference, I suspected the
absence of copper in the ore to be the cause of the fouling of
the solution, and after an experiment on a small scale had proved
the supposition to be correct, cupric chloride was prepared by
boiling blue vitriol with salt. One precipitation tank was cleaned
and filled with stock solution. Then 12 gallons of the prepared
cupric chloride solution was added, stirred and precipitated with
calcium sulphide. After precipitation the solution of this tank
was capable of dissolving nearly 50 per cent, more silver chloride
182 HYDROMETALLURGY OF SILVER
than before. This was tried in the laboratory with freshly pre-
pared silver chloride. The clear solution was decanted from the
copper precipitate and conveyed to the pump tank to be mixed
with the balance of the stock solution. During the first two
days the copper precipitations were made daily. A change in
the action of the solution could be observed at once. In all
nine copper precipitations were made, with a consumption of
127 Ib. of bluestone and 181 Ib. of salt. At the end of the ninth
precipitation the stock solution had almost regained its former
dissolving energy. The silver leaching time dropped to fourteen
and sixteen hours, and the value of the precipitate improved
again to $10 and $11 per pound. The solution kept in this
excellent condition for over a month, when signs of degeneration
could be observed; a fresh precipitation, however, brought it
right again. While not being prepared to offer a chemical expla-
nation why the want of copper in the ore should cause a degen-
eration of the sodium hyposulphite solution, and why an addition
of cupric chloride to this solution, which is precipitated before it is
sent in circulation, should restore to the solution its former quality
and dissolving energy, I place my observation on record with the
belief that it will be of great service to many lixiviation works.
In Cusihuiriachic, Mexico, it took seventy-two hours to leach
the silver from a tank charge of 8 tons, while after a treatment
of the stock solution with cupric chloride it was accomplished in
twelve to eighteen hours.
Filters for Leaching Tanks. — A free filtration being of great
importance for the success of the process, the selection of a proper
filter is, therefore, also of great importance. Some use gravel
and sand without a wooden filter bottom, as at La Baranca,
Sonora, Mexico. The material of such a filter is cheap enough,
but its preparation is troublesome, and therefore a filter is usually
kept in use until the filtration becomes so bad that it has to be
replaced by a new one. At Sombrerete straw filters were once
used. Short straw was spread on a wooden filter bottom about
a foot thick and then the ore was dumped on top of it. This
rather coarse filter did not produce a very clear filtrate when
new, but after the first charge it improved in this respect and
produced a clear solution; however, the outflowing stream de-
creased with every charge, and when the filter had to be renewed
it was found that the straw had rotted and packed tightly.
LIXIVIATION WITH SODIUM HYPOSULPHITE 183
Filters like those described above, which cannot be cleaned,
and are therefore kept in use as long as possible, invariably cause
a decrease of the working capacity of the leaching vats and with
it a decrease in production and percentage of extraction. Burlap,
the material used for making grain sacks, makes the best and
most lasting filter. It is cut in pieces and spread over the wooden
filter bottom so that one strip laps over the other about 3 in.
The ends are rolled up and packed tightly around the circum-
ference, or better into a groove as shown at N, Fig. 31. After
each charge the strips of burlap are carefully rolled up, so that
no residue which adheres to them may drop below the filter
bottom, are well washed and again spread on the filter bottom.
If this is done regularly the filter is kept in the best possible con-,
dition. If the washing is not done after every charge a thin
layer of residue will form on the filter cloth, gradually growing
thicker and harder, partly by the repeated pressure of the feet of
the laborers and partly by the deposition of calcium sulphate.
The gypsum also incrusts the fibers of the cloth, makes it stiff
and hard, and finally stops filtration entirely. Where burlap
cannot be obtained coarse sheeting is a good substitute.
End of the Silver Leaching. — The extraction of the silver from
a charge may be finished while the outflowing solution still gives
a considerable precipitate upon addition of calcium or sodium
sulphide. It is impossible to judge by the color of this precipitate
whether it contains silver or not. It is, however, important to
know when all soluble silver is extracted, in order to avoid unneces-
sary consumption of the precipitant and loss of time. To ascer-
tain when the end of the silver leaching has been reached a large
beaker is filled with the solution and some calcium sulphide is
added. The precipitate is allowed to settle, the clear solution
decanted, and the precipitate poured on a paper filter and washed
well. It is then removed from the filter and dissolved in nitric
acid, filtered to remove the sulphur, and a drop or two of hydro-
chloric acid is added. If a white precipitate or only a cloudiness
is produced, which by dilution and boiling does not disappear,
there is still silver in the solution and the leaching has to be
continued. If the solution remains clear the extraction is con-
cluded.
The influx of sodium hyposulphite is then stopped and the
solution is allowed to drain until it commences to disappear
184 HYDROMETALLURGY OF SILVER
under the surface of the ore, when a stream of water is turned on
to displace the sodium hyposulphite solution absorbed by the
ore. It is only necessary to continue this second application of
water for a comparatively short time; just long enough to keep
the same volume of stock solution on hand. Then the charge is
allowed to drain as dry as circumstances permit, after which the
residues are discharged. This is done by shoveling them into
chutes, of which one is placed between each two tanks and which
discharge into cars running underneath the tanks (Fig. 31), or
into a large triangular trough beneath the vats, in which a current
of water flows. The trough ought not to have an inclination less
than 1 in. to the foot. In some works the residues are sluiced
out. To make this method successful the vats must not be too
deep nor of too large diameter. The stream of water has to be
applied under pressure by means of a pump, or under good head
from a storage tank.
XIV
PRECIPITATION OF SILVER
IN all modern lixiviation works the precipitation vats are
provided with a mechanical contrivance to agitate the solution
during and after precipitation, and only in antiquated works is
this operation done by hand with a paddle. A horizontal beam
about 12 in. shorter than the inside diameter of the vat is fastened
to a vertical iron shaft (Figs. 57 and 58). This horizontal beam,
which moves above the surface of the solution, is provided with
hard-wood staves about 2 in. square and reaching down to about
1J in. above the bottom. These staves are so arranged that
when the agitator is in motion they cut the liquid with the edge.
The agitator is set in motion by a friction clutch, and it should be
started gradually to avoid breaking the staves. The inside of
the tank is provided with four vertical wooden wings, projecting
3 in. toward the center and reaching nearly to the bottom. They
break the violent current around the periphery and throw the
solution toward the center, thus causing a strong whirling motion.
The agitator should make 30 r.p.m. if the diameter of the vat
is not more than 8 or 9 ft. Agitators of this construction are
only durable when used in vats not deeper than 6 ft. In deeper
tanks the staves will break, and an agitator of stronger construction
has to be used.
Another method of agitation which is very convenient is by
the use of compressed air, furnished by a small compressor. From
the receiver a pipe-line leads along the whole row of precipitation
vats. At each vat is inserted a T with a f-in. branch. These
branch tubes are provided with valves and connected with a
rubber hose 6 to 8 ft. long, the other end of which terminates
with a f-in. gas pipe. This pipe must be long enough to reach
any point of the bottom and still project about 2 ft. above the
rim of the tank. If the vat is full and the air is turned on the
solution is put into violent motion, as if boiling. It is well to
185
186 HYDROMETALLURGY OF SILVER
change the position of the pipe from time to time, especially if
the vat has a large diameter. This method of agitating the
solution is now extensively used.
Preparing Calcium Poly sulphide. — This precipitant is pre-
pared by boiling milk of lime with pulverized sulphur. The
proportion has to be ascertained by the operator, as it depends
on the quality of the burnt lime rock. No free sulphur should
be visible in the white residues. It is soon found out how much
of the local lime has to be slacked for a certain weight of sulphur.
The boiling was formerly done in upright boilers, made of
boiler iron 4 ft. in diameter and 6 to 7 ft. deep, by direct applica-
tion of a steam jet. Near the bottom was a pipe outlet closed
with a wooden plug. This outlet discharged into an iron
tank in which a sand filter was prepared. Below the level of the
bottom of this filter tank was placed the iron storage tank to
receive the calcium sulphide solution. From this storage tank a
pipe-line conveyed the chemical to the precipitation tanks. The
boiling tank is first charged with the milk of lime, but enough
margin should be left for the water condensed from the steam,
which in the beginning is quite considerable. When the milk is
boiling the sulphur is charged gradually, by means of a sieve, in
order to scatter the same as finely as possible over the surface.
If charged by a shovel the sulphur sinks quickly to the bottom,
forming chunks which require a long time to combine with the
lime. When boiling is concluded some cold water is splashed
over the surface, which makes the foam disappear and also causes
a quicker settling. When settled the clear yellow-brown solution
is decanted, by means of a stiff 1-in. rubber hose, into the storage
tank. This done, the plug near the bottom of the boiler is re-
moved, and the sediment allowed to flow into the filter tank.
To make it flow more freely, some water is added before removing'
the plug. The filtrate from the residues also flows into the storage
tank. These residues are very slow in filtering and retain much
calcium sulphide solution, which has to be displaced by applying
water on top of the residues, which, however, is rather difficult
to do properly, as the residues are very pasty and very slow
filtering, so that, when they are removed by shovels, some of the
calcium sulphide will be wasted.
The calcium sulphide department was much improved lately
by myself. Fig. 35 represents a vertical section of the arrange-
PRECIPITATION OF SILVER
187
188
HYDROMETALLURGY OF SILVER
ment. A is the lime-slacking box, nearly level with the cool-
ing floor. When the lime is slacked the milk is allowed to flow
into the distributing trough B, whence it fills the boiler C.
There are two of these boilers, side by side, which can be alter-
nately charged through the distributing trough. This trough is
made of J-in. steel and is represented by Fig. 36. The outlets
in the bottom can be closed by wooden plugs.
The boiler C is represented in detail by Fig. 37. The sides
are made of f-in. while the bottom and top are of f-in. steel.
It is a boiler and pressure tank combined. Fig. 38 gives a top
view. There are six openings in the top, of which one is a man-
hole, while the other five are smaller, and are connected with
~~l t
-
--0—0-
"*
FIG. 36. — DISTRIBUTING TROUGH FOR MILK OF LIME.
To be made of steel.
pipes. Three of them reach nearly to the bottom (Fig. 35), viz.:
the discharge pipe passing through the center, the steampipe,
and the compressed-air pipe. The filling and air-escape pipes do
not extend into the boiler. The reason why the compressed-air
pipe reaches nearly to the bottom is to agitate the pulp during
discharging. The manhole cover has a 4-in. flanged opening in
the center (Fig. 39), which can easily be closed by a cast-iron
plate and bolts. This opening serves for introducing the sulphur;
it is left open during filling and boiling, and is closed during
discharging.
When the milk of lime is charged steam is turned on. The
opening in the manhole cover is kept open for the escape of the
PRECIPITATION OF SILVER
189
'16
FIG. 37. —BOILER AND PRESSURE TANK FOR
CALCIUM SULPHIDE.
Vertical section.
190
HYDROMETALLURGY OF SILVER
air during filling and for the escape of vapor during boiling.
Room has to be left for the water from condensed steam, and also
for foaming, which sometimes occurs when too much steam
FIG. 38.— BOILER AND PRESSURE TANK FOR
CALCIUM SULPHIDE.
Plan of head.
is admitted. By means of a wooden staff in which notches are
cut, and which is inserted through the opening in the manhole
cover, the filling is regulated.
It is not advisable, for two reasons, to make too strong a
FIG. 39. — BOILER AND PRESSURE TANK
FOR CALCIUM SULPHIDE.
Saddle to be riveted to head of tank.
solution. In a very strong solution crystals of bisulphide of cal-
cium are formed, which will be found on the sides and bottom of
the storage tank, and will also be formed in the pipe-line leading
PRECIPITATION OF SILVER 191
from that tank to the precipitation tank, and before long will
clog the pipe, necessitating the taking down of the whole line
for cleaning. The second inconvenience caused by too strong
a solution is the difficulty to see the end reaction in precipitating,
in consequence of which the precipitator is apt to add too much
of the precipitant.
I found that a solution answers well if, in boiling, to
each cubic foot of water about 2J Ib. of sulphur are taken.
Of good and freshly burned lime about 2i Ib. to the pound
of sulphur is an average proportion, but, as stated above, it
depends entirely on the quality of the lime.
If it is intended to fill the boiler three-quarters full, the cubic
content is calculated and by it the amount of required sulphur
and lime ascertained. The lime is then slacked and the milk
charged into the boiler, then water is added to fill the boiler
three-quarters full. Then steam is admitted. The sulphur is
charged gradually when the water commences to boil. Boiling
has to be continued for four or five hours, according to circum-
stances. If in a sample, taken with a long-handled iron ladle from
the bottom after filtering, some free sulphur should be found,
boiling should be continued. If this has no effect it shows that
lime is wanting, which has to be added. In the next charge
more lime has to be taken in the beginning.
When boiling is concluded the steam is turned off and the
volume of the solution measured with the wooden staff; if the
volume is short, it has to be filled up with water to the mark in
the staff. Then the opening in the manhole cover is closed tight
and compressed air is applied. The discharge pipe is connected
with a 24-in. filter press, as shown in Fig. 35. The filtrate flows
into the calcium sulphide storage tank, and from there is conveyed
along and above the precipitation tanks, which are also shown in
this figure. When the filter press is filled with residues, water
under pressure is forced into the press to wash the residues.
The wash-water is allowed to run into the storage tank and mix
with the strong solution. In the floor near the lime-box is a
chute, discharging into iron cars. The residues from the lime-
box as well as from the filter press are thrown into this chute.
Adding the Precipitant. — The precipitant, whether sodium or
calcium sulphide, is kept in a reservoir made of boiler iron, from
which it is conveyed through an iron pipe to the precipitation
192 HYDROMETALLURGY OF SILVER
vats. At each vat there is attached to the pipe a hose, which
is closed by a pinch-cock. In commencing to precipitate it is
well to open the hose a little and to throw, by swinging it, some
of the precipitant over the surface. By the appearance of the
clouds which are formed, whether heavy or light, the precipitator
can see at once if he has to precipitate a concentrated or dilute
charge. Then he sets the solution in agitation and allows the
precipitant to flow in. An experienced precipitator can judge
by the color which is created when he splashes some of the pre-
cipitant over the surface, while the solution is agitated, the
progress of precipitation, and knows when it is nearing the end.
While precipitation is going on the clouds which are formed
become gradually lighter in color, and toward the end almost
yellow. When nearly finished the influx of the precipitant is
stopped, and after a few minutes the agitation also. Then the
flaky precipitate is allowed to sink somewhat below the surface,
and some of the precipitant is splashed over the surface. Accord-
ing to the appearance of the clouds more or less precipitant is
added and the solution is agitated again. This operation is re-
peated until precipitation is complete. If by an addition of the
precipitant no reaction takes place, it is well to throw some strong
silver solution over the surface after the precipitate has partially
settled. If the places where the silver solution fell turn reddish
brown, the precipitant is in excess and. more silver solution has
to be added. One who is not experienced would best make this
test in a beaker.
The precipitation tank as illustrated in Fig. 35 is 12 ft. in
diameter and 8 ft. deep. The solution is agitated by compressed
air. The two outlet pipes, P and N, are of lead. To the upper
pipe, P, is attached on the inside of the tank the decanting hose,
M, with the float, S. This should be a very stiff hose. It is kept
above the solution, while the tank is filling or precipitation is
going on, by a thin rope fastened to a hook on the outside of the
tank. When the precipitate has settled, the rope is unhooked
and the end of the hose allowed to float. The float is so arranged
that it keeps the end of the hose immersed. When the float comes
in near approach to the precipitate close attention has to be
paid, so that no precipitate is carried out. As the decanted solu-
tion is to be used again for extracting silver, it is conveyed by
means of troughs, running along in front of the precipitation
PRECIPITATION OF SILVER 193
tanks or through a pipe-line to the lower storage tanks, in which
it is collected and elevated to the upper storage tanks, which are
placed on a higher level than the rim of the leaching tanks. The
outer end of the pipe P is connected with a short piece of hose,
which lies in the trough, if troughs are used, to avoid splashing.
The lower outlet pipe, N, is provided with a valve and con-
nected with an iron pipe-line, common to all precipitation tanks,
which leads to a pressure tank, by means of which the solution
is forced into a filter press. The upper hose P1 serves for ad-
mitting the precipitant.
XV
TREATMENT OF THE PRECIPITATE
The Precipitate. — In precipitating the base-metal solution
we have seen that not all the metals present are equally affected
by the sulphur of the precipitant, and that the silver especially
is more readily precipitated than the other metals. This is also
the case if these metals are dissolved in the sodium hyposulphite
solution, and therefore the precipitate which is obtained in the
earliest stage of precipitation contains far more silver than that
obtained later. Thus the operator has it in his power to make
different grades of precipitate. This, however, does not offer
such advantages as in base-metal leaching, and is of no direct
practical value, because in order to maintain the dissolving energy
of the sodium hyposulphite solution it is absolutely necessary to
precipitate as perfectly as possible all the metals dissolved in it;
but it explains why the black layer of sulphides, which we fre-
quently find deposited on the surface of the ore charge, is so
much poorer in silver than the precipitate itself. If the precipi-
tation was done well, and ample time was given to the precipitate
to settle, and the decantation of the solution was always performed
properly and without mishap, we should not find any black deposit
on the top of the ore charge; but such exact work is not always
done, especially in large works, which seldom have a sufficient
number of vats to give the precipitate ample time to settle.
The different sulphides settle according to their respective
specific gravities. Lead, silver, and copper go down first, while
antimony, zinc, iron, and free sulphur follow. While this sepa-
ration is not theoretically perfect, it takes place to such a degree
that the particles which settle last may contain but 30 to 100 oz.
silver per ton, while the total precipitate may contain 5000 to
15,000 oz. per ton.
This black layer of sulphides deposited on top of the ore charge
194
TREATMENT OF THE PRECIPITATE 195
after an extended lixiviation, being so much poorer than the
precipitate, does not involve any notable loss of silver if it is
carefully scraped off before the residues are discharged and the
scrapings are mixed with the ore and roasted.
F. Sustersic's Method of Preparing the Precipitate for Refining
by Cupellation. — If a precipitate contains a large percentage of
copper, the refining of the same by cupellation with lead requires
a large amount of lead, causes the formation of a large amount
of rich by-products, and increases the cost of refining.
For the treatment of such a precipitate F. Sustersic devised
a method by which he extracts the copper first, leaving the
precipitate in an excellent condition for cupellation. Reporting
on very interesting experiments made with such precipitate, he
says:
"The crude precipitate produced by lixiviating the ores of
Avino, Durango, Mexico, with sodium hyposulphite was analyzed
and shown to contain:
Silver 4.206 per cent.
Gold . . 0.2064 per cent.
Copper 14.80 per cent.
Lead 16.40 per cent.
Iron • . . . 0.70 per cent.
Zinc 0.50 per cent.
Chlorine 5.10 per cent.
Lime 3.87 per cent.
Sulphur 39.40 per cent.
Insoluble 4.20 per cent.
Not ascertained 10.80 per cent.
"This precipitate is rather base, which was caused by the
fact that the ore is very susceptible to heat and readily loses
considerable silver by volatilization at only a moderate roasting
temperature, and therefore had to be roasted at an extremely
low heat, at whicji practically all base-metal salts remained in
the roasted ore. To refine such a precipitate with sulphuric acid
would be too expensive in Avino, and the refining with lead by
cupellation would offer considerable difficulties on account of the
large percentage of copper, so it became necessary to find a
proper method by which the copper could be removed from the
precipitate before the latter is subjected to cupellation.
" In roasting a material containing copper sulphide and silver
sulphide the copper is converted into sulphate at a very low
heat, while it takes a bright-red heat to convert the silver into
196 HYDROMETALLURGY OF SILVER
sulphate.1 In accordance with this difference in the property of
these two metals I made several tests and obtained very gratifying
results.
" A six-inch roasting dish containing 100 grams of the precipi-
tate was placed into a muffle the temperature of which was kept
below red heat. As soon as the free sulphur commenced to burn
the roasting dish was removed from the muffle and the charge
continually stirred until the sulphur flame ceased. Then the
charge was again placed into the muffle and roasted at so low a
heat that the chemically combined sulphur oxidized without
ignition. This second operation required only fifteen minutes.
The color of the precipitate changed from black to greenish gray.
Weight of raw charge 100.00 grams.
Weight of roasted charge 85.40 grams.
Loss in weight 14.60 per cent.
The roasted precipitate ought to assay 4.925 per cent, silver.
By actual assay it was found to contain iJ*28() per cent, silver.
The loss by volatilization, therefore, was Nil.
"An analysis of the roasted precipitate gave the following
result :
Silver 4.9286 per cent.
Copper 18.50 per cent.
Lead 19.30 per cent.
Ferric oxide 1.75 per cent.
Zinc 0.30 per cent.
Lime 8.60 per cent.
Sulphuric acid 37.70 per cent.
Insolubles 6.80 per cent.
"Ten grams of the roasted precipitate were leached with
water until the filtrate became colorless. No trace of silver
could be detected in the filtrate. The copper in solution was
precipitated with zinc.
The 10 grams of roasted precipitate contained 1.85 grams copper.
The nitrate was found to contain in solution 1 .66 grams copper.
Copper extracted as sulphate 89.73 per cent.
" The residues on the filter weighed 5.8 grams, and the roasted
precipitate therefore lost 42 per cent, of its weight by removing
the soluble substances with water. These residues were analyzed
and found to contain:
" * See Chapters on " Sulphating Roasting," and "Ziervogel's process."
TREATMENT OF THE PRECIPITATE 197
Silver 8.50 per cent.
Copper 1 .60 per cent.
Lead 33.50 per cent.
Iron 1.65 per cent.
Lime 6.00 per cent.
Sulphuric acid 28.50 per cent.
Taking the loss in weight into calculation which the roasted precipitate sus-
tained by leaching with water, if there was no silver dissolved the resi-
dues should contain 8.4965 per cent, silver.
By actual assay it was found that they did contain. . . . 8.5000 per cent. silver.
Therefore, silver dissolved by leaching with water was . . Nil.
"These results show that by this method nearly 90 per cent,
of the copper in the precipitate can be removed and recovered
as metallic copper if the roasting is properly executed, while all
the silver and the lead remains concentrated in the residues,
which are then in a state well suited for cupellation.
" By several check-tests it was proved that no loss of silver
is sustained, neither in roasting the precipitate nor in leaching
the same with water."
This method of F. Sustersic is undoubtedly a very rational
way to prepare the precipitate for cupellation. Not only does
it remove and recover nearly 90 per cent, of objectionable copper,
but also it removes other salts, thus reducing the weight of the
roasted precipitate by 42 per cent., which is connected with a
corresponding enrichment in silver and lead. It permits the
production of a clean litharge and greatly reduces the formation
of rich slag or froth. There is nothing to diminish the usefulness
of this method if the free sulphur of the crude precipitate is re-
moved by boiling with caustic soda instead of by burning. On a
commercial scale the precipitation of the copper from the solu-
tion is, of course, done with scrap iron.
Accumulation of Sodium Sulphate in the Solution. — I stated
above that when the leaching with water is stopped and the charge
is ready for silver leaching the ore still contains some sodium
sulphate, which during silver leaching will enter the stock solu-
tion. In course of time the sodium sulphate will therefore accu-
mulate to such a degree in the stock solution that the latter will
greatly lose in dissolving energy, a much longer leaching time will
be required, and the percentage of extraction will suffer much.
When such trouble arises operators usually try to "freshen up"
the solution by adding more sodium hyposulphite to it. This,
however, will benefit it only for a very short time, and the same
198 HYDROMETALLURGY OF SILVER
trouble will appear again. Some add more hyposulphite every
day, thus increasing greatly the cost of extraction without get-
ting the solution back to its original energetic condition.
Calcium Sulphide as Precipitant. — The best method is to
conduct the process so that no sodium sulphate will accumulate
in the stock solution. This can be done by using calcium
sulphide as precipitant. Sodium sulphate reacts with calcium
sulphide, forming sodium sulphide and calcium sulphate, which
precipitates, while the sodium sulphide acts as precipitant for
the metal chlorides dissolved in the solution. On account of
this reaction, which is of so great an advantage to the lixiviation
process, I always advocate the use of calcium sulphide as pre-
cipitant.
The valuable effect of this reaction is especially felt if ore is
treated which requires a large percentage of salt in roasting,
whereby larger quantities of sodium sulphate are produced.
Should so much sodium sulphate enter the stock solution that by
the regular process of precipitation all the sodium sulphate is not
decomposed, this salt will then gradually increase, notwithstand-
ing the use of calcium sulphide, and will spoil the solution. It is
therefore advisable to add at certain intervals an excess of cal-
cium sulphide, and finish precipitation with silver solution.
This will free the solution entirely from sodium sulphate.
If sodium sulphide is used this reaction does not take place,
and not only will all the sodium sulphate dissolved during silver
leaching accumulate and remain in the stock solution, but the
amount will be increased from the precipitant, which always
contains more or less sodium sulphate. In such works it is well
to make provision for the manufacture of calcium sulphide, solely
for the purpose of purifying the stock solution from time to time.
Treatment of the Precipitate. — The precipitate should be dis-
charged from the precipitation tanks every other day. If it is
allowed to accumulate for a longer time it changes from a flaky
condition to a very fine powder, which settles slowly and will
cause the solution in circulation to contain in suspension a con-
siderable amount of exceedingly fine precipitate. The precipi-
tate is discharged into a tank with a slowly moving agitator,
varying in size according to the size of the works, from which the
precipitate is drawn into a pressure tank in portions as required,
and thence, by means of compressed air, it is forced into a
TREATMENT OF THE PRECIPITATE 199
Johnson filter press. Pressure tanks made of cast-iron will resist
longer the corroding action of the solution than those made of
boiler iron. They should be made so that the top or cover can
be removed when required. The relief pipe should return to
the sulphide tank connected with the agitator, because the air
when relieved escapes with great force, and is apt to carry along
some precipitate. Instead of compressed air, steam may be used;
but the pressure will be limited by the pressure in the boiler, and
often may not be sufficient. The feed pipe of the press should be
in connection with water and air pipes, so that after the filter
press is filled the precipitate can be washed by pumping warm
water through it, and partially dried by forcing compressed air
through it. When so treated the precipitate will come out of
the press in hard cakes, permitting a clean handling of them.
The press should be placed in a separate room with cement floor.
Such a floor is easily kept clean, and prevents loss of silver.
Instead of a pressure tank, a steam pump can also be used;
only attention should be paid that the steam cylinder of the pump
is larger than the pump cylinder, so that a pressure of 150 Ib. or
more can be produced with it, if it should be necessary. The
pump cylinder should be provided with a bronze piston rod and
liner, because iron is more or less affected by the solution, and in
course of time the inside of an iron cylinder becomes very rough.
Much cleaner work can be done with a pressure tank, because
there is always more or less leakage around the piston rod of a
pump. The filtrate is conveyed to one of the lower solution
reservoirs.
It was found more convenient, cleaner, and labor-saving not
to use an agitating tank for the precipitate, but to connect the
pressure tank direct with the pipe-line which runs in front of the
precipitation tanks and with which they are in communication.
The filling of the pressure tank is done direct from the precipi-
tation tanks, but in order to release them as soon as possible for
their regular work, the pressure tank is made large enough to
receive the whole of the precipitate of one tank. It is made of
steel and used in an upright position.
When the charge has been forced into the filter press and the
pressure tank is empty, the compressed air contained therein is
allowed to escape by opening the valve of the air-escape pipe.
The air rushes out with great force, carrying out moisture and
200
HYDROMETALLURGY OF SILVER
some of the precipitate hanging on the side near the air outlet,
which, if left to escape free, would not only cause some loss of
silver, but make the surroundings very unclean. To avoid this
I designed and introduced a drum into which the air is discharged,
and which gives very good satisfaction. Fig. 40 is a vertical
section. The drum is 24 in. in diameter and 3 ft. 6 in. high. In
alternating distances of 6 in., 4 in., 6 in., 4 in., etc., angle irons
A, A, are riveted to the inside, forming circular shelves. On
these shelves rest conical trays, one with the cone turned down,
FIG. 40.— AIR BLOW-OFF DRUM.
Vertical section.
the other with the cone turned up. Those with the cone down
have a 6-in. circular opening in the center, while the others have
openings (C, C, Fig. 41) near the periphery. These trays are
kept in place by four small bolts. Underneath the bottom tray
is an outlet pipe leading to one of the lower storage tanks, while at
an angle of 90 deg., enters the air-escape pipe from the pressure
tank. The working of this air blow-off drum is clearly explained
by the drawing.
The Pressure Tanks. — If a pressure tank is destined to lift
liquors containing residues or a precipitate it ought to be used
in an upright position, because it facilitates the discharge of these
TREATMENT OF THE PRECIPITATE
201
solids, but if such a tank is used only for clear solutions, then it
is of more advantage to have it in a horizontal position. Fig. 42
represents a horizontal pressure tank designed to lift the sodium
FIG. 41. —AIR BLOW-OFF DRUM.
Plan of tray.
hyposulphite solution. It is made of steel 12 ft. long and 4 ft.
6 in. in diameter. The filling pipe P enters at the head of the
tank in order to save grade. This tank has to be placed in a pit
below the bottom level of the lower storage tank; by inserting
the filling pipe at the center of the head a more shallow pit will
Steel
FIG. 42.— HORIZONTAL PRESSURE TANK, FOR SOLUTION.
answer, and is more convenient for work. The manhole is on
top, but near one end of the tank. At A the solution discharge-
pipe is inserted, which extends nearly to the bottom. The part
202
HYDROMETALLURGY OF SILVER
of the pipe which extends inside the tank is subject to wear and
should be so arranged as to permit of quick replacement. In
Fig. 43 its construction is shown in detail: A, a short tube, one
FIG. 43. — CAST-IRON FLANGE UNION FOR DISCHARGE PIPE OF
PRESSURE TANK.
The dimensions here given are for 2-in. discharge pipe as used in the vertical
pressure tanks. They have to be changed for other pressure tanks according to
size of their respective discharge pipe. The part that extends into the tank can
be easily removed and replaced.
end of which is fastened to the tank, while the other is provided
with a flange F, of the shape shown. B, the pipe which extends
into the tank. The upper end of it is provided with the flange
TREATMENT OF THE PRECIPITATE 203
E, which is faced on both sides. This flange fits loosely into
the recess of the flange F. When the tube is inserted a J-in.
rubber gasket, C, is placed on top and covered with the flange L,
which is attached to the discharge-pipe M. This done, the
flanges L and F are tightly drawn together by the bolts D, D.
The rubber gasket is strongly pressed by the bolts and makes a
perfectly tight joint. By the construction it can be seen that
the pipe B can be withdrawn and replaced in a few minutes. All
pipes extending into a pressure tank ought to be arranged in the
way just described.
Handling of the Sodium Hyposulphite Solution. — It was
stated above that the solution, by precipitating the silver and
the other metals dissolved in it, is regenerated, and can be used
over and over again indefinitely. As the solution works from
the upper level down to the lowest level of the leaching plant by
gravity, it has to be elevated again in order to keep it in circula-
tion. The solution coming from the precipitation tanks is col-
lected in a number of large but not too deep tanks (storage tanks).
These tanks, usually four in number, communicate by means
of iron pipes near the bottom. The solution coming from the
precipitation tanks flows only into the first tank and into no
other, but through the communicating pipes all four tanks are
filled simultaneously. The fourth or last tank only is connected
with the filling-pipe of the pressure tank. This arrangement of
the four communicating tanks offers an excellent opportunity
for settling any precipitate that may have been drawn out in
decanting by carelessness. Above the leaching tanks there is
another set of four storage tanks of the same dimensions, arranged
exactly as are the lower tanks. Into the first of these tanks the
pressure tanks lift and discharge the solution, while from the
fourth one a pipe-line leads over all the leaching tanks. These
upper storage tanks give another opportunity for settling. They
are cleaned once or twice a year.
In large works the circulating stream of solution is quite
voluminous, and it is advisable to have two solution pressure
tanks, so that a constant stream can be maintained.
Removal of Sulphur. — The precipitant being a polysulphide,
the precipitate will contain a large percentage of free sulphur,
whether calcium or sodium sulphide is used. It is desirable to
remove this free sulphur from the precipitate before the latter
204
HYDROMETALLURGY OF SILVER
is subjected to a final treatment. The best method is to boil
it with caustic soda in an iron tank, the caustic soda combining
with the free sulphur and forming sodium sulphide, which serves
as precipitant for the silver; but care has to be taken that no
excess of caustic soda enters the stock solution, because it will
exercise a decomposing action on the silver chloride in the ore.
To avoid this the sodium sulphide solution thus obtained is con-
veyed to the boiling vessels in which the precipitant is manufac-
tured. After boiling with caustic soda the precipitate shrinks
FIG. 44. —APPARATUS FOR THE MANUFACTURE OF LYE.
To be placed on cooling floor next to roaster floor.
much in volume and becomes very heavy. To separate it from
the sodium sulphide solution after the main part has been de-
canted is a very slow process if done by common filters, and
therefore it is much better to use a small filter press for this pur-
pose. In works where no filter press is used, and a common
filter has to be employed, these badly filtering sulphides can be
made quick filtering by treating them with a strong silver solu-
tion to decompose all the sodium sulphide, which is the cause of
the bad filtration. By this method 60 per cent, of the sulphur
contained in the precipitate can be regained and brought into
TREATMENT OF THE PRECIPITATE 205
a state in which it can be used again as precipitant, thus greatly
reducing the actual consumption of sulphur.
F. Sustersic proposed to leach wood ashes, convert the lye
by boiling with caustic lime into caustic potash and boil the pre-
cipitate with it, thus producing potassium sulphide, which is
used as precipitant. I adopted and carried out this method.
In most works wood is used as fuel for the roasters and boilers,
and the ashes are thrown away. By adopting this method a
large part of these ashes can be utilized. Fig. 44 illustrates an
arrangement for the manufacture of lye. The tank T is placed
near the roaster floor, with its rim just a little below that floor,
so that the wheelbarrow runway resting on the rim of the tank
is on a level with the roaster floor. The tank is provided with a
filter bottom (not shown in the figure). The ashes from the
roasters and also from the boilers, if the latter are conveniently
situated, are wheeled and dumped into the tank, which is filled
one-third full with water. This is done to protect the tank and
filter from the hot ashes, which contain many small pieces of burn-
ing coal. When the tank is filled, water is allowed to flow on top
of the ashes. The outflowing lye is collected in the iron tank R.
When the lye begins to get weak the hose H is placed in the
trough L, which leads outside the building, and the charge is
allowed to drain and then replaced by fresh ashes. From the
iron tank R a pipe-line leads to the calcium sulphide boilers,
which can also be used for the manufacture of caustic potash.
One of the boilers is charged with lye, to which milk of lime is
added. The mixture is then boiled. Samples are taken from
time to time, filtered, and to the filtrate a few drops of hydro-
chloric acid added. If this causes effervescence boiling is con-
tinued, but if after a while effervescence is still caused by the
acid, then more milk of lime has to be added. W^hen finished
the charge is pressed through the same filter press which is used
for calcium sulphide. The filtrate is collected in an iron storage
tank placed alongside of the calcium sulphide tank.
Fig. 45 represents a system of three pressure tanks; they
serve for the treatment of the precipitate, and are placed in the
refinery of the works. The pressure tank A is charged with
precipitate from the precipitation tanks through the pipe D.
From there it is forced through the pipe E into a filter press.
When pressed, the precipitate is charged into pressure tank C,
206
HYDROMETALLURGY OF SILVER
through the opening in the manhole. Caustic potash solution
is added, which is done by opening the valve of pipe G, which
latter is connected with the caustic potash storage tank. The
mixture is boiled, and when finished forced, through pipe H, into
a second filter press. The filtrate from this press, which is potas-
sium sulphide, flows into pressure tank B, through pipe K, whence
it is lifted up through the pipe M and charged into the calcium
sulphide boiling tank.
FIG. 45. —PRESSURE TANKS FOR TREATMENT OF
PRECIPITATE.
As by this operation about 60 per cent, of the sulphur is re-
gained in a shape in which it can be directly applied as precipitant,
and at the same time the precipitate is freed from its surplus sul-
phur, and as, besides, the caustic potash does not cost more than
the labor and the lime, this is surely a very economical operation.
TREATMENT OF THE PRECIPITATE 207
Another way of expelling and regaining the free sulphur
from the precipitate is by distillation. The moist precipitate is
charged into retorts and heated, the sulphur vapors being con-
ducted to brick chambers and condensed as flowers of sulphur.
I used this method years ago on a large scale with satisfactory
results as to the amount of sulphur regained, but the cast-iron
retorts did not last long enough, principally on account of care-
lessness on the part of the men in charge, who overheated them;
and as transportation of such heavy castings into the mountains
was very difficult and expensive, this method was discarded.
However, with careful firing, and in localities where transporta-
tion facilities are better, it can be applied to great advantage.
While the extraction of the free sulphur will not be so complete as
by boiling the precipitate with caustic soda or potash, the opera-
tions are fewer. The product is dry and ready immediately for
further treatment, while in the former method the product has
to undergo the processes of filtering, washing and drying.
A third method, by which, however, the sulphur is lost, con-
sists in burning it off in a small reverberatory furnace. An actual
roasting is not required; in fact it ought to be avoided, to prevent
loss by volatilization. The sulphides ought to be charged dry,
but if they are charged moist they should remain undisturbed in
the moderately heated furnace until dry, to avoid the generation
of rapidly evolving steam, which is apt to carry away fine par-
ticles of the already dry part of the precipitate, thus causing a
loss. Even if the precipitate has been previously dried in special
ovens, the heat in the beginning has to be kept very low for some
time, to avoid mechanical loss by steam, because only seldom
will the precipitate be perfectly free from moisture after leav-
ing the drying ovens. Later the temperature is increased to ignite
the sulphides. They commence to burn with a blue flame near
the fire-bridge, and the flame spreads gradually over the whole
charge. When this takes place the fire has to be lowered to
avoid overheating. No stirring should be done until the flame
ceases; then a gentle stirring is given. This brings up new flames,
which, however, do not last long, but reappear if the charge is
stirred again. Stirring is repeated until the flame ceases entirely.
The temperature has to be kept so that when the flame ceases
the charge is perfectly dark, which indicates that no actual roast-
ing of the material took place and that only the free sulphur was
208
HYDROMETALLURGY OF SILVER
burned off. If the precipitate is treated in this way no loss by
volatilization will take place, and that is all which is required,
because for the further treatment of the precipitate an actual
roasting, a changing of the sulphides to oxides and sulphates, is
not necessary; in fact, is hurtful. Mr. Stetefeldt, in his book on
lixiviation, stated the loss to be as high as 6 and 12 per cent., but
this is not so.1 However, there is a chance of loss by this method
if it is not properly executed, especially if the precipitate con-
tains antimony or is charged wet into the furnace.
T1
FIG. 46. — DRYING AND ROASTING FURNACE FOR SILVER
PRECIPITATE.
Fig. 46 represents a vertical and Fig. 47 a horizontal section
of a small reverberatory furnace for burning the silver precipitate.
Where the boiling of the precipitate with caustic potash or soda
is adopted this furnace can be used for drying the treated pre-
cipitate.
The first method, i.e., boiling with caustic soda or potash, is
the most rational and entirely excludes any loss of silver, and
offers the additional advantage of regaining about 60 per cent,
of the sulphur.
Refining the Precipitate. — The refining is done with litharge
1 Stetefeldt based his statements on reports received from Cusihuiriachic,
while the works were under the management of Mr. Dagget. Later it
developed that the great loss attributed to the burning of the sulphides was
caused by the dishonesty of the man in charge of that part of the process,
who -stole systematically part of each charge.
TREATMENT OF THE PRECIPITATE
209
on a lead bath in the cupeling furnace. Other methods have
been tried, but so far not with much success.
If the refining is done on the lead bath usually English cupel-
ing furnaces are used, which are so constructed that the test can
be dipped toward the front to pour the refined silver into the
molds. If, however, the works produce larger quantities of precip-
itate, it is more advantageous to cupel in a larger furnace, simi-
lar in construction to the German cupeling furnace, with the
exception that the bottom is not stamped direct into the circular
space left in the brickwork, but into a circular cast-iron pan, the
bottom of which is perforated with J-in. holes. The pan is
FIG. 47.— DRYING AND ROASTING FURNACE FOR SILVER
PRECIPITATE.
made in four sections, which are kept together with a few bolts,
which construction permits the pan to expand without breaking.
This pan sets above a second pan, 2 in. larger in diameter
and only 2 in. deep, and rests on bricks set in the lower pan on
their 4-in. side. This flat pan serves as a guard against loss of
silver, in case it should happen that the bottom of the test cracks.
The space underneath communicates with the outside of the fur-
nace by four channels, through which air circulates and cools
the bottom, while in drying the test they serve as vents for the
vapors. The perforated pan is 5J to 6 ft. in diameter and 12 in.
deep, and has a square cut 12 in. wide and 10 in. deep for the
litharge bridge. This cut is placed toward the front. To the
210 HYDROMETALLURGY OF SILVER
left of it is the fireplace and to the right the flue, which is pro-
vided with a damper, while the back of the furnace is arranged
to admit the blast. The furnace proper is covered with a dome
made of boiler iron and lined with clay. This dome is lifted and
can be swung to one side by a crane.
A cheap and excellent material for the cupel is a mixture of
one part of clay (by volume) and three parts of lime rock, pul-
verized and well mixed. Both ingredients must be free from
quartz and ore particles, for which reason, if the crushing has to
be done by machinery that is also used for pulverizing ore, such
machinery must first be cleaned very carefully.
After the material is well mixed part of it is spread on a clean
floor, sprinkled with water, and quickly worked with shovels so
that the mixture becomes uniformly moist. The mixture should
not be made too moist, as otherwise in refining the bottom is apt
to come up. The material is in proper condition if a handful of
it, squeezed hard, forms a ball, which may be handled gently, but
should crumble into its former condition by a slight pressure with
the fingers. In preparing the test the perforated pan is filled
about 6 in. with the prepared material, then stamped down with
iron bars.
These bars are made of a piece of round iron 1 J in. in diameter
and 8 to 10 in. long, one end of which has the shape of an egg
while the other is welded to a IJ-in. gas-pipe 5 ft. long. The
material in the pan is leveled and beaten in by two men stand-
ing on boards laid across the brickwork. They commence to
beat in the center, pursuing a spiral course, stamping with the
egg point in a perpendicular direction, by a lift of about eight
inches, striking with the rod close to each preceding stroke.
When by a screw-like advance the stamping has reached the side
of the pan, it has to be carried on back to the center in the same
way, then again to the side and so on, till about two inches of
loose material remains.
If a hole can be scratched easily with the finger in the
stamped mass, the bar must be used with more force. Care
must be taken always to have still two inches of loose material
above the stamped mass when a new charge is put in, because
if the whole is beaten hard, the next charge will not unite per-
fectly with the under layer. The beating on the second charge
is done in the same way, and so on until the hard-beaten mass
TREATMENT OF THE PRECIPITATE 211
reaches about two inches below the rim of the pan. Then a
6-in. high iron ring fitting the rim of the pan is placed on the rim, a
new charge added, and stamping continued until the hard stamped
mass reaches about three inches above the rim of the pan.
Then the loose material is removed, the ring pulled up, and the
hard mass is cut down roughly by means of a hatchet or an adze
to about an inch above the rim of the pan, and then leveled with
a sharp short-handled scraper. This done, a circle is scratched
in, leaving a margin of about four inches around the periphery
of the pan. Inside this circle the hard stamped mass is cut out
spherically with a short-handled adze about 7 to 8 in. deep; then
made smooth by a scraper with a curved blade. When the
bottom is finished the hood is put over it by means of the crane
and a gentle fire started for drying. It is well to scatter a thin
layer of ashes over the surface. The fire should be kept about
twelve hours, then the ashes removed and the lead charged.
When the lead is melted and becomes red hot some coarsely
pulverized litharge is charged, enough to cover the whole surface
of the bath. A strong fire is kept and the precipitate, mixed with
litharge, is put on the bath in small charges. When this is going
on the draft has to be checked by the damper so that the flames
will come out through the working door. This precaution is
taken to reduce the amount of such valuable dust to be carried
out by the draft into the dust-chambers. After the charge is
introduced the damper is opened again, and when the charge is
melted the upper layer is worked gently with a hoe, and a new
charge is introduced. This is repeated until all the precipitate is
charged. Then some more litharge is added, the working door
closed and the strong fire continued. It will be noticed that,
when everything is well melted, around the periphery of the
bath many small bubbles will appear. This is caused by the
carbonic acid which the lime rock gives off, and does not injure
the bottom.
When the last charge is well melted the top layer is drawn off
over the litharge bridge by means of a hoe until the surface be-
comes bright. Then the blast is turned on, playing on the sur-
face, which makes the bath fume profusely. This is caused by
the oxidation of the lead matte which had formed during melting
and which lays on top of the lead. These fumes smell strongly
of sulphurous acid. After two or three hours the surface loses
212
HYDROMETALLURGY OF SILVER
...A.
TREATMENT OF THE PRECIPITATE 213
its brightness, a yellow-red ring of litharge forms around the
periphery, and little islands of the same color are formed where
the blast strikes. They float about until they reach the ring
of litharge around the periphery and unite with it. The width of
this ring increases until nearly the whole surface of the bath is
covered. When this has happened a flat channel is cut into the
litharge bridge, through which the litharge flows off. The flow
is to be regulated so that the ring is kept about 8 in. wide. Care
is to be taken that no metal flows over the bridge. The tempera-
ture has to be kept so that the litharge runs freely over the bridge
but cuts the channel very little. If the heat is too high the
litharge will cut the channel rapidly deeper and metal will flow
out. On the other hand, if the temperature is too low the litharge
will flow sluggishly over the bridge and form a soft crust on the
inside, near the channel. Too low a temperature has to be
carefully avoided, as the whole bath may freeze, especially to-
ward the end of the process. Likewise should a too high tem-
perature be avoided, because then the litharge cuts around the
periphery of the bottom, thus ruining it, and eats out the channel
too quickly, besides which more silver and lead are volatilized.
At the end of the operation, when less litharge is formed,
the bath becomes covered with a net-like coat, moving on the
convex surface and consisting of litharge, between which the
silver glances through in spots. These spots grow larger, till at
last the net breaks and the litharge slides to the sides, producing
a display of colors. Before the end, but at the time when the
formation of litharge becomes scanty, the iron molds are arranged
on an iron bench set in front of the furnace. A long-handled iron
ladle hooked to a chain reaching down from above is heated by
placing it to one side inside the furnace. As soon as the surface
of the silver turns bright it is ladled into the molds. During
ladling the temperature has to be kept high.
Should the litharge become stiff and sluggish toward the end,
even if the temperature is increased, it is a sign that there is not
enough lead in the bath to separate the copper, and without loss
of time some lead has to be added, but on the side of the test
and not into the bath, because this would cool it too much and
might freeze it.
All the products which are formed in cupeling contain silver.
The first slag drawn, as well as the last litharge near the end of
214 HYDROMETALLURGY OF SILVER
the operation, is the richest, while the poorest and purest is
obtained during the middle of the operation. The latter goes
back to the process at the next cupellation. The products that
are too rich it is best to sell to the smelting works, as it would be
too expensive to treat them by themselves.
As it cannot be avoided that some dust of the precipitate,
which is very valuable, is carried away by the draft during charg-
ing, and as the lead fumes from the bath also contain silver, it is
very necessary to provide for an effective dust collector. Fig. 48
illustrates a four-shaft system of O. Hofmann's dust collector.
It is inserted in the main flue of the refinery, so that all the fumes
and dust from the different furnaces have to pass through it.
XVI
CONSTRUCTION OF TROUGHS
In lixiviating works, troughs are extensively used for con-
veying the base metal and the silver solution, and it is of great
importance to have and keep them perfectly tight. The solu-
tions will corrode metals, including lead, and the troughs there-
fore have to be made of wood, but it is a very difficult task to
make and keep troughs perfectly tight, especially if they have
only a slight inclination and the solutions do not move swiftly.
Leakage occurs principally in places where the boards are spliced.
White lead, tar, pitch, putty, etc., are of but little avail. White
lead is the worst, and ought never to be used by a carpenter to
make a joint waterproof. As long as the paint is wet it will
be tight, but when the wood absorbs the oil and the paint dries,
it contracts into numerous threads and wrinkles and hardens,
forming minute channels through which the solution will find
its way. The places where the white lead was applied will not
swell, and the trough will be much less tight than if no cement
had been used at all. In order to convey the solution to the
different tanks, branch troughs have to be used. One end of
these troughs is placed under the main trough, and right above
it one or more holes are bored into the bottom of the main trough.
These holes are closed by long wooden plugs. The loosening and
tightening of these plugs is done by mallets, and therefore the
main trough is subjected to considerable rough usage, and the
pitch or putty will crack and start leakage.
In running the large blue vitriol plant of the Consolidated
Kansas City Smelting and Refining Company of Argentine,
Kansas, which I designed and erected, I was very much annoyed
by trough leakage. There were over 2000 ft. of troughs in use.
To guard against leakage all the troughs were lined with 6-lb.
215
216 HYDROMETALLURGY OF SILVER
sheet lead properly put in by experienced lead burners, but be-
fore a year passed the troughs began to leak in different places.
Searching for the reason it was found that this leakage was
caused by a peculiar property of the lead. The copper solution
passing through the troughs was hot. It did not flow continually,
but in charges, so that the lead was exposed alternately to the
hot liquor and to the cooling effect of the air. The lead lining,
which originally was perfectly smooth, was found to be full of
wrinkles. These wrinkles could have been originated only by the
expansion of the lead. By special experiments I convinced my-
self that lead is a metal which when heated expands, but when
cooled does not contract as much as it had expanded; and if a
sheet of lead is alternately heated and cooled it continues to grow
larger and larger. If there is no room for free expansion, as in a
trough or tank, the sheet has to fold up in wrinkles. If this is
continued, the sheet gets so thin in certain places that it finally
breaks.
For some time the leakage was stopped by burning a piece
of sheet lead over the leaking places, but before long the leaks
became so frequent that the item of trough repairing became
seriously high, and it was concluded to renew all the troughs.
Lead lining was of course discarded. Previous experiments
with different kinds of soft wood showed that California redwood
resisted best and longest the action of a hot copper solution.
Other experiments were made to prepare a cement which softened
but did not melt at a temperature of 90 to 96 deg. C., and which
in cooling did not become hard and brittle but remained pliable.
Such a cement was found by boiling lard oil with rosin and rubber
and red oxide of iron. Waste pieces of sheet rubber and old
pieces of rubber belting were used to supply the rubber. The
fibers of the belting were removed with iron hooks and forks.
The oxide of iron was added last. The boiling was done in an
iron kettle with a fireplace below. The cement or paste was
applied hot. Fig. 49 shows the construction of the trough in
cross-section. The joints of sides and bottom are made step-
shaped, as shown in the drawing, and are first coated with the
hot cement, then screwed tight with brass screws. The corners
are filled with triangular wooden moldings, which are first well
coated with cement. They are kept in place and tightly drawn up
by brass screws. If the corners are first well cemented, covered
CONSTRUCTION OF TROUGHS
217
by a strip of thin sheet rubber, and then the molding screwed on,
additional security for tight joints is obtained. The troughs are
made in sections of the length of the boards, and both ends cut
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FIG. 51. -TROUGH UNION.
square. The ends of two sections are brought together, a pure
rubber gasket, made to order in one piece, J or f in. thick, is
218 HYDROMETALLURGY OF SILVER
placed between them and drawn tight by bolts. Fig. 50, a side
view, illustrates the manner in which the two ends of the trough
are drawn together. C and D are flanged iron castings of the
shape of the cross-section of the trough, but smaller, so that they
will fit into a groove 2J in. wide and i in. deep cut into the three
sides of the trough and about 6 in. from the end. These castings
are made in pairs — one right, one left, as shown in Fig. 51. When
placed, they are fastened to the trough with wood screws, and
by means of the bolts E the ends of two troughs are pressed to-
gether. Thus perfectly tight joints of the different sections are
obtained. At distances of four feet wooden pieces are nailed
across the trough to prevent the sides from spreading.
Troughs made in this way are perfectly tight and will stay
so for years.
XVII
TROUGH LIXIVIATION
IN tank lixiviation, the extraction of the silver from chlori-
dized ore by solutions of hyposulphite salts is performed by fil-
tration. The ore particles are kept stationary, while the solvent
moves down through the mass of ore. The quickness of extrac-
tion, other conditions alike, is in direct proportion to the rapidity
of the movement of the solvent through the ore. The solution,
if left in contact with the ore without moving, displays but very
little dissolving energy. If the filtration is interrupted for ten or
twelve hours, and thus solution and ore are left in complete con-
tact for that length of time, it will be found that, when filtration
is started again, the outflowing solution is but very little more
saturated with silver than it was at the time of interruption,
and that the ten or twelve hours were almost a complete loss in
the total time of extraction. Notwithstanding the long contact,
the solution had not become saturated with metal chlorides to
its full dissolving capacity. A rapid movement of the solvent
through the ore is essential to a quick extraction. This fact is
well known; and the endeavor of leachers has been to hasten
extraction by increasing the rate of filtration. Siphons, vacuum
pumps and other devices have been used with more or less success,
but none of them has given full satisfaction.
I have found that, if chloridized ore, after the base-metal
chlorides are removed, is brought into rapid contact with a
proper volume of moving sodium hyposulphite solution, the
silver chloride contained in the ore dissolves almost instantly,
and that it is rather the volume of the solvent than its concen-
trated state which produces this effect. Such favorable con-
ditions cannot be attained in tanks. The rapidity with which
a certain volume of the solvent can be brought into contact
with the ore particles is limited by the speed with which the
219
220 HYDROMETALLURGY OF SILVER
solution descends through the ore; and thus the leaching time
in tanks cannot be shortened beyond the limit set by the filter-
ing capacity of the ore. In a trough, however, these favorable
conditions can be attained by gradually introducing the ore into
the moving stream of the solvent. The ore can thus be brought
into rapid contact with any desired quantity of the solvent, and
moves in and with the stream. The effect is astonishing. Ore
charged at the upper end of a trough not longer than 12 to 15 ft.
will leave the trough at the lower end as tailings, having yielded
all its silver chloride to the solvent. This is accomplished dur-
ing the very short time of 4.7 seconds which it takes the pulp to
rush through the trough. But in order to obtain satisfactory
results, that is, to extract all the silver chloride contained in the
ore as shown by the chlorination test assay, it is necessary, and
of great importance, to maintain a certain proportion of solvent
and ore, which proportion depends on the nature of the ore. By
numerous experiments I have found that all kinds of silver ores,
no matter how differently they behave in tank lixiviation with
regard to the length of time required for the extraction of the
silver, will yield their silver chloride in the same short time.
They behave all alike in this respect, but only if the proper pro-
portion between solvent and ore which each respective ore re-
quires is maintained. It is interesting to observe that, when two
ores of different chemical character but of equal filtering property
are treated in the trough, the one which in tank lixiviation re-
quires the longer time to yield its silver chloride to the solvent
will need in the trough a larger volume of solvent than the ore
which requires in the tank less time for the extraction of the
silver.
Lead-bearing ores require a long leaching time, for the reason
that lead sulphate reduces greatly the dissolving energy of sodium
hyposulphite for silver. In the ordinary lixiviation the solution
becomes more saturated with lead sulphate as it descends through
the ore and loses proportionally its dissolving energy. As the
solubility of the lead sulphate increases with the concentration
of the sodium hyposulphite solution, a stronger solution does not
hasten the process; but if we bring the ore rapidly in contact
with a large volume of hyposulphite solution, the latter retains
enough of its dissolving energy to produce a quick silver extrac-
tion. The presence of lead sulphate, therefore, does not retard
TROUGH LIXIVIATION 221
trough lixiviation; it merely entails the use of larger quantities
of solvent.
Results of much interest, obtained by me in experimenting
with ore from the Cusihuiriachic Mining Company, Chihuahua,
Mexico, illustrate the importance of maintaining a certain pro-
portion of solution and ore to obtain satisfactory results. The
ore contains considerable lead, and the extraction by common
tank lixiviation required on an average nine hours for base-
metal and 53.8 hours for silver leaching, in all 62.8 hours. It
was roasted in Howell furnaces. For this experiment a 1.6 per
cent, solution was used. The roasted ore was first leached with
water to remove the base-metal chlorides before treating it in
the trough.
The roasted ore contained 27.12 oz. silver per ton.
The chlorination test tailings called for . 3.94 oz. silver per ton.
The ore and solution passed through 43 feet of trough.
PROPORTION OF SOLUTION AND ORE IN WEIGHTS
6 weights solution to 1 ore, the trough tailings contained
14.58 oz. silver per ton.
8 weights solution to 1 ore, the trough tailings contained
6.56 oz. silver per ton.
12 weights solution to 1 ore, the trough tailings contained
5.25 oz. silver per ton.
18 weights solution to 1 ore, the trough tailings contained
4.37 oz. silver per ton.
24 weights solution to 1 ore, the trough tailings contained
4.37 oz. silver per ton.
As the length of the troughs was in all cases alike, the time
of contact of solution and ore was the same, and as the strength
of the solution was also similar, the difference in the extraction
was caused only by the volume of solution. The results show
that the extraction is in direct proportion to the volume of solu-
tion until the maximum is reached, when an increase of solution
does not improve the extraction any more. The results further-
more show that for the Cusihuiriachic ore the proportion is 18
weights of solution to one of ore, probably less, but more than
12 to 1. In this experiment the trough tailings contained 0.43 oz.
per ton more silver than the chlorination test tailings called for.
This shortage in extraction was caused by the fact that the 1.6
per cent, solution was too strong, as numerous subsequent ex-
periments demonstrated. A solution containing but 0.5 per cent,
sodium hyposulphite gives, as in tank lixiviation, the best results.
222 HYDROMETALLURGY OF SILVER
The idea suggested itself to investigate the behavior of the
base-metal chlorides contained in the roasted ore in a moving
stream of water, and it was found that they dissolved just as
rapidly as the silver chloride in the sodium hyposulphite solution,
and that passing through the same short length of trough the
solution was accomplished, with the exception of a certain per-
centage of sodium sulphate, which, however, the ore is always
found to contain in tank lixiviation at the time when base-metal
leaching is stopped and silver leaching commenced.
We have learned in tank lixiviation that, by leaching the ore
charge with water in order to remove the soluble metal chlorides,
the water becomes so charged with these chlorides that it dissolves
silver chloride like sodium chloride does, and that the solubility
of the silver chloride increases with the concentration of the solu-
tion. To investigate this important feature of the lixiviation
process, I made the following laboratory test: Freshly prepared
silver chloride was introduced into solutions of sodium chloride
which had a temperature of 120 deg. F., and was left in contact
for five minutes, during which time the solution was vigorously
agitated, and then filtered. The filtrate was tested for silver
with sodium polysulphide, and it was found:
Solution of 2 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes no silver.
Solution of 3 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes no silver.
Solution of 4 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes no silver.
Solution of 5 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes no silver.
Solution of 6 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes no silver.
Solution of 7 per cent, sodium chloride at 120 deg. F. dissolved in
5 minutes some silver.
The filtrate of the 7 per cent, solution showed a faint coloring
by sodium sulphide. When using an 8 per cent, solution the
coloring was still very faint. Molten silver chloride poured
into a 5 per cent, sodium chloride solution at 100 deg. F. decrep-
itated into fine powder; but the filtrate did not show any reac-
tion for silver.
The result of this experiment shows that, if it were possible
to regulate the base-metal leaching in tanks so that the outflow-
ing solution at no time contains more than 6 per cent, of metal
chlorides, the same would not dissolve any silver and could be
TROUGH LIXIVIATION 223
allowed to run to waste, instead of being subjected to a special
treatment to regain the dissolved silver.
It is different in trough lixiviation. There we have it in our
power to regulate the grade of concentration of the resulting
base-metal solution at will, and we can therefore produce at once
a solution which is sufficiently dilute not to dissolve silver chloride.
This is one of the great advantages of trough lixiviation. Should
the ore contain sufficient cupric chloride to make the saving of
the copper an object, the whole resulting solution may be
passed through a series of tanks or deep troughs filled with scrap
iron. Its diluted state will not prevent the chemical reaction.
I have made a very interesting observation in respect to the
solubility of the base-metal chlorides with regard to the trough
principle. Two samples of complex ores from different mining
districts were roasted with 8 per cent, of salt. Twenty grams of
each were placed on a paper filter and leached with water. When
all the soluble base-metal chlorides had been extracted, the fil-
trate was weighed in both instances, and it showed that the
amount of water required for sample No. 1 was three times the
weight of the ore, while for sample No. 2 it was seventeen times
the weight of the ore. By treating the same quantity of ore one
and a half minutes on the trough principle, no perfect extraction
of the base metals could be obtained, until the same proportion
of water and ore was used as that required in common leaching,
viz.: 3 to 1 for ore No. 1, and 17 to 1 for ore No. 2. Both these
ores were rather base. This behavior of the soluble chlorides
toward the water as solvent is undoubtedly remarkable. If
for both ores we assume the filtering property to be alike, it would
follow, according to the quantity of water required, that if leached
in tanks ore No. 2 will have to be leached 5§ times as long as ore
No. 1 to extract the soluble salts. The experiment, however,
showed that this extraction can be accomplished from both ores
in the same short time by bringing them at once in contact
with their respectively required quantity of solvent.
Care has to be exercised in taking samples in trough lixivia-
tion. The final residue sample can be taken from the tank with
sampling irons in the usual way before the residues are dis-
charged, but samples required for the observation of the process
during operation have to be taken in a different way. The whole
stream has to be caught in a vessel of proper size, say in an enam-
224 HYDROMETALLURGY OF SILVER
eled iron kettle of two or three gallons capacity. The proper
place to take the sample is where the stream leaves the trough
and drops into the settling-tank. The kettle is quickly pushed
under the stream so that it receives the whole stream, and quickly
withdrawn as soon as it is filled. With the proper mark
attached to it the kettle is left undisturbed until the liquid be-
comes perfectly clear, then it is carefully decanted and filled up
again with water and stirred well. This washing is done twice,
when, finally, the settled pulp is evaporated to dryness, well mixed
and quartered down.
In the laboratory, experiments can be made on the principle
of trough lixiviation by introducing 20 grams of roasted ore,
which, if the test is for silver, has previously been washed, into a
graduated cylinder of 1000 c.c., in which is contained 100 c.c.
or 200 c.c. of sodium hyposulphite solution, according to the
proportion which is intended to be used. The top of the cylin-
der has to be tightly closed with the palm of the hand, and the
cylinder has to be brought into a horizontal position, and then
oscillated in order to make the ore and solution pass quickly from
one end to the other, to imitate the current in a trough. This
is done for about three-quarters of a minute or for one minute,
then the contents of the cylinder are poured into a filter, washed
with water to displace the silver solution from the sand and paper,
dried and assayed. The same operations are required in experi-
menting with the base-metal chlorides, except that water is used
instead of sodium hyposulphite solution.
Having by experiment found the required proportion of
water and ore and solution and ore, the size of pumps, pipes,
outlets, etc., can be calculated.
THE TROUGHS
A triangular shape of the troughs is preferable, because the
same quantity of water will display more energy for moving the
sand than on a flat bottom. An inclination of three-fourths of
an inch per foot is sufficient. We have seen that when the pulp
passes through a trough 12 to 15 ft. long the extraction is ,com-
pleted, therefore no particular attention has to be paid to the
length of the troughs. The necessary trough connection from
feeder to sluice-tank and from there to the settling-tank give
more than enough length. The stream in the trough moves
TROUGH LIXIVIATION 225
swiftly, and therefore very little bottom pressure will be exerted,
so that it is very easy to construct these troughs so that they will
be tight.
SLUICE-TANKS AND SLUICING
As the ore has to be moved from the tanks by a stream of
water, it is not advantageous to give the tanks a large diameter.
Twelve or 14 ft. is sufficient, though if circumstances demand it
larger tanks can be used.
Figure 52 represents the vertical section, Fig. 53 the horizontal
view, and Fig. 54 the front view of a settling-tank arranged for
sluicing. In the center of the bottom is the discharge opening,
6 in. in diameter. The cast-iron discharge-tube, k, of the same
inside diameter, tightly fastened to the outside of the tank bottom,
corresponds with the discharge-hole. The lower end of the tube
is at right angles to the upper end, and provided with flange o.
The valve ra, which is provided with a rubber gasket, can be
pressed tightly against flange o by turning the wheel F. Flange
o and valve ra are made of brass. Fig. 55 shows in detail the
construction of wheel F. Part of the valve-stem is square and
rests at ra in a square box, so that by turning the wheel F the
valve m does not turn too, but moves forward or backward. By
this arrangement the life of the rubber gasket is much lengthened,
as no turning force is exercised against the flange o, but only a
quiet pressure. Around the discharge opening, and fastened
to the bottom of the tank, is the wooden polygon v, in which is
cut the groove pr Around the inner periphery of the tank,
and high enough to give the filter bottom an inclination of at
least three-quarters of an inch to the foot, is the groove p. Fig. 53
illustrates the construction of the filter bottom, which is made
in sections. The filter cloth is well fastened, and kept in place
by driving tightly a rope into the grooves pt and p. The air-
escape pipe d, which reaches to the rim of the tank, enters the
latter close under the filter bottom. A piece of hose is fastened
to the upper end and can be closed by a hose clamp. In Fig. 54
#!, q, are solution outlets; s, filter outlet. Connecting pipes g and
h (Figs. 52 and 54) have, like the discharge-tube k, to be well
coated with asphaltum varnish. In the same figures z is the
water-pipe; n, the central hose, which ought to be very stiff to
resist the pressure of the ore; it reaches down into the discharge-
226
HYDROMETALLURGY OF SILVER
SETTLING VAT •- HORIZONTAL VIEW- Scale;- Jfct-1 ft.
FIGS. 52 and 53. — SETTLING-TANK ARRANGED FOR SLUICING.
TROUGH LIXIVIATION
227
tube k, where it has .to remain during the process of charging.
Before charging the tank the discharge-tube is filled with water
through the central hose, in order to keep the latter filled with
water, which will prevent the inside of the hose from being ob-
structed by ore. In the base-metal department the central hose
has to be connected with both the water- and solution-pipe.
The connection with the solution-pipe serves for sluicing the ore,
FIG. 54.— SETTLING-TANK, FRONT VIEW.
Scale, i in. = 1 ft.
while the water connection is used after the tank is empty to free
the tank and filter cloth from all adhering sodium hyposulphite
solution by rinsing, otherwise it would get into the base-metal
solution and dissolve some silver. In the silver department
only the connection with the water-pipe is required, as the
hose is used only for sluicing out the residues.
When a tank is ready to be discharged, the wheel F is turned,
and thus the valve m pulled back. The water is injected through
228
HYDROMETALLURGY OF SILVER
the central hose, while the latter is gently -moved up and down.
The stream undermines the tightly packed sand, causes a con-
tinual caving in, until a funnel-shaped opening is made through
its depth to the surface. Then several streams are made to play
on the top, while the central hose, with checked stream, is left
in position to avoid obstruction of the discharge-tube by a too
sudden rush of sand.
The central position of the discharge opening and the funnel
shape of the filter permit a quick and clean sluicing. The pulp
leaving the discharge-tube enters the sluice-trough t underneath
the tank, which leads to the tailings-trough u in the silver depart-
FIG. 55. — WHEEL FOR CLOSING DISCHARGE GATE.
Scale, i in. = 1 in.
ment, or to the silver-leach trough in the base-metal department.
The charge being sluiced out, tank and filter have to be well
rinsed; the valve of branch pipe a is opened, and the adhering
sand washed off from flange o and valve ra by the double sprinkler
w. During discharge the valve m has to be pulled back far
enough to prevent the outflowing pulp from striking it, otherwise
the rubber gasket would soon wear out. Then the discharge
valve is closed again, and the tank is ready to be connected with
the other tanks. In Figs. 52 and 54 y represents the silver-
TROUGH LIXIVIATION 229
leach trough; b, the intersecting box above the tank, and i a hose
made of duck.
ARRANGEMENT AND OPERATIONS
The construction and manipulation of the sluicing tank was
treated previously to the description of the general arrangement
and operations of a trough lixiviation plant in order to make it
better understood.
Solution is performed outside the tanks in troughs, while the
ore is moving in and with the stream of solvent, and the tanks
are used only to separate the solids from the liquid. The system
is a continuous one; but as the lixiviation process requires two
solvents, first, water for the removal of the base-metal chlorides,
and then a solution of sodium hyposulphite for the extraction of
silver, it has to be divided into two departments, the base-metal
and the silver departments. Fig. 56 shows a complete arrange-
ment. The upper series of tanks represents the base-metal, the
lower series the silver department. The tanks in each depart-
ment are placed on the same level and close together. They are
connected by pipes a, 6, c, d, e, and /, in such a way as to form a
perfect circuit. These connecting pipes are placed a few inches
below the rim of the tanks and also on a level. The diameter of
these pipes depends on the daily capacity of the works and the
proportion of solvent and ore to be used. Each tank has two
outlet pipes qv q, Fig. 54, on a level with the communicating pipes,
and one S, from under the filter. They all discharge into the
base-metal solution trough which leads outside the building,
either to waste or to scrap-iron tanks for the recovery of copper.
Each of these tanks is arranged and constructed for sluicing, as
described above.
The silver-leaching department consists of the same number
of tanks of the same size and construction as those of the base-
metal department, but they are placed on a lower level. They are
also connected with communicating pipes forming a circuit.
On a still lower level are placed the precipitation tanks, which
in this case are agitated by mechanical stirrers, and on the next
platform we find the filters for the precipitate. Below this are
the solution-pump vats.
The triangular trough extends above, and in the center
line of each row, and wherever the trough branches off, there is
230
HYDROMETALLURGY OF SILVER
FIG. 56. — SYSTEM FOR CONTINUOUS TROUGH LIXIVIATION.
For a working capacity of 50 tons per day.
TROUGH LIXIVIATION 231
inserted an intersecting box which, if located above a tank, is
provided with a plug-hole in the bottom, through which the pulp
flows if that tank is to be filled. The troughs branching off from
the intersecting boxes are each provided with a sliding gate,
which is kept closed except in that trough into which the stream
is to be directed.
The operations are as follows: The roasted ore from the cool-
ing-floor is charged into an iron hopper which at its bottom has
an adjustable worm discharge. The worm discharges into a
stream of water flowing into a triangular trough which feeds the
pulp into a grinding machine, in which the lumps are mashed,
and which will be explained further on. Passing this lump-
mashing machine, the pulp enters the triangular base-metal
leach-trough, and is thus conveyed automatically to the base-
metal tanks. The first tank to be filled is tank C, Fig. 56; but
before commencing operation, the following preparations have
to be made. The slide-gate of the trough leading from the
intersecting box to tank C is opened, while those of the other
two troughs are closed. The connecting pipe c, between tanks C
and G, is closed by a plug from the inside of C. The outlets of
all the tanks are closed except the outlets ql q, near the rim of
tank G, Fig. 54. Likewise all outlets from under the filters are
closed. Then the central hose n, Fig. 52, is inserted into the dis-
charge-tube, the valve of which is also closed. This done, the
pulp is permitted to flow into tank C, Fig. 56. When the solution
reaches the level of the connecting pipes, it will flow through
connecting pipe b into tank B, and when B is filled into tank A,
and so on until the last tank G is filled, when the solution will
leave the tank through the two outlets near the rim. As soon
as tank C is filled with ore, the pulp is made to flow into tank B.
The connecting pipe 6 is closed, and thus tank C is disconnected
from the circuit. The outlet s, below the filter of tank C, is
opened, and the solution is allowed to drain into the base-metal
solution trough. When the solution begins to disappear below
the surface of the ore, water is admitted, to press out the solution
absorbed by the ore. When this is done, sodium hyposulphite
solution is applied to press out the water. As soon as the liquid
flowing out through filter outlet s shows indications of silver, the
outlet s is closed, and the charge is ready to be sluiced for silver
leaching. While tank-charge C is under the described treatment,
232 HYDROMETALLURGY OF SILVER
which does not take much time, tank B is gradually filling with
ore. When filled, the pulp is made to enter tank A, and tank B
is disconnected from the circuit, and its charge is treated in the
same way as that of C.
When the charge C has been prepared as described above,
it is sluiced out with sodium hyposulphite solution to extract the
silver. Underneath the tank the pulp enters the silver leach-
trough and flows down to tank C, of the silver department. The
arrangement and construction of the tanks are the same as in
the base-metal department, and when operations have been
started the connections have to be set the same as described for
base-metal leaching.
Charge C, being sluiced out with solution, the filter-outlet s
is turned into the silver leach-trough, the hose clamp is opened,
and the solution contained under the filter is allowed to flow
out, and then the inside of the tank and the filter is rinsed with
water, which is also allowed to flow into the silver leach-trough.
This accomplished, outlet hose s is closed and returned to the
base-metal solution trough. Then the plug of connecting pipe
c is removed, and tank C is restored again to the circuit. C being
empty, the flow in the base-metal solution trough will cease until
the tank is filled again with base-metal solution.
By using the proper proportion of solvent and ore, the latter
will drop as tailings into the tank, while all the silver chloride
which can be extracted by the common chlorination test will be
dissolved by the solution. It is well to have a special solution
pipe by which a stream can be made to flow directly into the
silver leach-trough close to the tank, so that if the volume of
solution used in sluicing should not be sufficient, the proportion
can be maintained by the use of this extra pipe.
The clear solution leaving the last tank Gf flows through the
silver solution trough into the distributing-box, and from there
to any desired precipitating tank. The bottom of the distribut-
ing-box has twice as many holes as there are precipitating
tanks. In these holes bent lead pipes are inserted from below
and are fastened by flanges. Stiff two-inch hose-pipes, of which
two lead to each precipitation tank, are attached to these pipes.
The holes can be closed by long plugs. This arrangement I
found quite convenient. The operator can direct the stream
from the main working floor, without being obliged to creep over
TROUGH LIXIVIATION 233
all the precipitating tanks, as is customary in many lixiviating
works.
When one tank is filled with tailings, it is disconnected from
the circuit and the pulp is admitted into the next tank. Outlet
sv under the filter, is opened; the solution still contained in the
tank is allowed to drain off, and the part retained by absorption
is displaced by water. The tailings are then sluiced with water.
Where water is scarce, the wash-water can be collected and used
for sluicing out the final tailings.
It is of advantage to connect the solution-pipe with the solu-
tion-pump. In the first place a higher pressure can be obtained
than from the storage tanks, and on the other hand by the num-
ber of pump-strokes per minute the volume of solution can be
calculated. The extra stream into the trough outside the tank,
if one should be used, it is better to have from the storage tank, as
the high pressure from the pump would splash in the trough.
PRECIPITATING VATS
Figs. 57 to 61 illustrate the construction of precipitating
vats and a convenient arrangement of filters of the precipitate
which I devised and have operated. This construction of tank
and filters is particularly adaptable for works which are not
equipped with air-compressor and filter-presses. The tanks are
provided with machine-stirrers, of the construction indicated
in the drawings. The stirrer s (Figs. 57 and 58) has to make
about 30 r.p.m. if the diameter of the tank is not more than
8 or 9 ft. It is set in motion or stopped by working the friction-
clutch / (Fig. 57). The wings, g (Fig. 58), which reach near to
the bottom, are about 3 in. wide and are kept in position by
triangular pieces of boards. They break the violent current
around the periphery and throw the solution toward the center,
thus causing a strong whirling motion. In Figs. 58 and 59, a is
the discharge-pipe, or decanter, for the clear solution, and b for
the precipitate. Pipe d, Fig. 59, passes in front of all the precipi-
tation tanks, and conveys the calcium or sodium sulphide solu-
tion from the reservoir to the tanks. The branch pipe c reaches
above the rim of the tank and ends in a rubber hose, which is
provided with a clamp. In precipitating, the stream can be
conveniently regulated by the use of the clamp, and the operator,
by observing the color produced by the precipitant in the moving
234
HYDROMETALLURGY OF SILVER
solution, can finish this operation in a very short time, and much
more easily than by using buckets.
One man can precipitate three tanks at a time without assist-
ance. The solution is so thoroughly agitated that a very perfect
separation of the silver sulphide takes place. The separation
is so perfect that the bottom of the tank can distinctly be
FIG. 57.— PRECIPITATION TANK, VERTICAL SECTION.
seen through 5 ft. of solution. To produce a quick and perfect
separation of the precipitated silver sulphide, the solution has
to be vigorously agitated. This cannot be well done in tanks of
14 or 16 ft. in diameter. It is much preferable to have smaller
tanks and a larger number of them. A good size is 8 to 9 ft. in
TROUGH LIXIVIATION
235
diameter and 6 ft. deep. In some leaching works where there
are large precipitation tanks in use, we find the bad practice of
discharging the precipitate only once a week, in some even only
once a month. Fresh precipitate forms large flakes, which settle
easily and cleanly. After two or three days it assumes a dry and
sandy condition, and if stirred up divides into very fine particles,
which are kept suspended in the solution for a long time. The
result of such a practice is that the decanted solution will not be
free from precipitate when used again for extraction, and a black
~~Scale 1A^ I ft.
FIG. 58. — PRECIPITATION TANK, PLAN.
coating of precipitate will cover the surface of the ore in the
leaching tanks. The precipitate ought therefore to be removed
from the precipitating tanks every day, and should never be
allowed to remain longer than two days. By using smaller
tanks with the machine-stirrer, and by discharging the precipitate
every day, or every other day, the circulating solution can be
kept so clear that, even after a prolonged leaching of five or six
days, no black coating can be observed on the top of the ore.
The machine-stirrer does good work also in discharging the
236
HYDROMETALLURGY OF SILVER
precipitate. After the clear solution has been decanted within
a few inches of the precipitate, the stirrer is set in motion, and the
discharge-pipe b opened. The stirrer agitates the precipitate
MACHINE
FOR HOFMANN'8
CONTINUOUS TROUGH LIXIVIATION
FRONT VIEW
FIG. 59. — PRECIPITATING VAT.
until nearly all is discharged into the niters, or into the pressure
tank by which the filter-press is filled. Thus the cleaning of a
tank can be done in a very short time.
TROUGH LIXIVIATION
237
Where the filter-press is not used, a filter arrangement as shown
in Figs. 59, 60 and 61 will be found very convenient. Two rows
of filters are so arranged that they are in communication with
each other by depressions cut into the divides of the frame. By
allowing the precipitate to flow into one filter, all the adjoining
filters will gradually be filled, one after the other, without requir-
ing the attention of the operator. The filters, which are made of
common cotton sheeting, are shallow, and the precipitate will
thus be spread in a comparatively thin layer over a large filter-
ing surface. Under each row of filters is placed a trough, n,
which receives the filtrate and conveys it to the pump-tank P,
below the floor. These filter frames are placed in front of the
precipitating tanks and can be made to contain quite a number
FIGS. 60 and 61.— FILTER FRAME.
Fig. 60 is a horizontal view. Fig. 61 is a section on line A-B.
of filters, as shown in Fig. 56, where two sections serve for five
precipitating tanks. Owing to the shallowness of the filters
and the large filtering surface, the solution will drain off fast,
and in five or six hours the precipitate will be stiff enough to be
charged with wooden hand-paddles into the drying furnace.
The extraction in trough lixiviation is not produced by fil-
tration; and the silver as well as the base-metal chlorides con-
tained in the lumps of the roasted ore will not be extracted if
the lumps are allowed to enter the trough together with the finer
portion. Actual grinding is not necessary, but in order to ob-
tain good results it is necessary first to mash the lumps. An
agitator will not perform the work rapidly enough, and some
quick-grinding machine has to be employed for this purpose.
238
HYDROMETALLURGY OF SILVER
Figs. 62, 63, 64 and 65 represent a grinding machine specially
designed by myself for use in trough lixiviation. The mantle m
is concave-shaped and is stationary. The muller /, a flatter cone
than the concave of the mantle, is inserted from below, and can
be lowered or raised by the screw g, for the purpose of regulating
the fineness to which the lumps are intended to be mashed.
FIG. 62. — LUMP-GRINDING MACHINE, ELEVATION.
Mantle and muller are provided with exchangeable shoes and dies.
The centrifugal force developed by the rotation of the muller
greatly assists the discharge of the pulp. Toward the center
the shoes and dies are provided with teeth, while toward the
periphery their surface is smooth (Figs. 64 and 65). The teeth
cut the larger lumps, the smaller ones are mashed by the smooth
part of the cones. Water and ore are charged through A, while
TROUGH LIXIVIATION
239
the circular cast-iron trough t receives the pulp and conveys it to
the base-metal leach-trough o. The canvas strip k prevents the
pulp from flying over the rim of the trough. By far the main
portion of the roasted ore is fine enough to pass through without
^i ft.
FIG. 63. — LUMP-GRINDING MACHINE, PLAN.
being affected, and only a small part will have to be mashed.
For this reason, and on account of the softness of the wet material,
large quantities of ore can be put through in twenty-four hours
with but very little wear of the machine, which in itself is of
simple and cheap construction.
240 HYDROMETALLURGY OF SILVER
PRACTICE OF TROUGH LIXIVIATION AT CUSIHUIRIACHIC.
The Don Enrique Mining Company at Cusihuiriachic had
accumulated a large dump of second-class ore, which contained
about 25 oz. silver per ton, but filtered so badly on account of the
large amount of porphyry in the gangue that it was not profitable
to work it by tank lixiviation. After the large lixiviation works
of this company were destroyed by fire, I undertook to work this
second-class ore by trough lixiviation, and treated 12,000 tons
by this method, at a good profit.
The work was done in the old North Mexican mill buildings,
which were formerly used for tank lixiviation until the ore in the
company's mine gave out. There were two sluicing-vats with
-i ft. ^--_r %-i ft.
FIG. 64. — Mantle. FIG. 65. — Muller.
LUMP-GRINDING MACHINE.
central discharge and flat, funnel-shaped filter bottoms, which
were placed about 10 ft. above the rim of jthe old leaching-vats, of
which there were eight, arranged in two rows. Each vat measured
14 ft. in diameter and 3J ft. in depth. These tanks were used as
settling-tanks, and were connected each with the other by a
4-in. pipe inserted near the rim, thus forming a circuit of all the
vats, so that if one tank was filled the solution could flow into
the other, and from there into the next, and so on. The filter
bottoms of these vats were covered with sheeting and a 4-in.
layer of washed river sand. On the cooling-floor a small hopper
with funnel-shaped bottom was erected and covered with an
inclined screen of J-in. mesh, to prevent the lumps from entering.
The hopper discharged to a short screw conveyor, and the con-
veyor into the cups of a belt elevator. The elevator lifted the
TROUGH LIXIVIATION 241
roasted ore and discharged it into a short triangular trough, in
which a stream of water was running. The trough was so
arranged that the pulp could be conveyed to either of the two
sluicing-tanks.
The speed of the screw conveyor being always the same, the
feed of roasted ore into the trough was uniform, and the desired
proportion of ore and water was easily regulated by the stream
of water. When the first vat was filled the pulp was made to
enter the second sluicing-vat. While the second vat filled, the
liquid in the first vat became clear and was drawn off by siphons,
so that when vat No. 2 was full vat No. 1 was ready to receive
again the stream of pulp. This was repeated until both vats
were fairly filled with washed ore. After the first charge of pulp
the outlet under the filter of each vat was opened, discharging a
clear stream while filling was in progress, the ore, which had by
this time settled on the bottom, acting as filter. This increased
the filling capacity of the vats. When both tanks were charged
the solution above the ore was allowed to drain, and then water
was applied to displace the solution absorbed by the ore.
The silver-leaching troughs were not longer than was necessary
to reach to each of the eight settling-vats. The intersecting
boxes over each settling-vat were 18 in. square and 12 in. deep,
with a 3-in. round opening in the bottom, which could be closed
with a wooden plug. To prevent splashing during charging a
4-in. canvas hose was fastened around the opening in the bottom,
reaching down to within a few inches below the rim of the settling-
vat.
Shortly before sluicing began the plug in the box above the
first tank was removed and a piece of filter cloth, kept in place
by several bricks, was spread directly under the canvas hose, to
protect the sand filter. The pulp, dropping always on the same
place in the vat, the ore, or rather the residues (because the silver
is already extracted when the pulp drops into the vat), will form
a cone, which, however, will never project much above the sur-
face of the solution, because the material in the solution, being
loose and lighter than the material above the solution, will slide
down. Thus a tank of 14 to 16 ft. diameter will be charged
pretty evenly. To fill the lower space around the periphery a
short trough was placed under the drop, and by changing the
position of the trough gradually the stream was directed to all
242 HYDROMETALLURGY OF SILVER
points. Long before the vat was filled with residues the solution
reached the level of the communicating pipe and flowed into the
next tank (No. 2), and then into No. 3 and No. 4. When tank
No. 2 was filled with solution the outlet under the sand filter
was opened and a clear stream of solution was discharged. The
same was done when No. 3 was filled with solution, and so on.
The filtration was very free, and the volume of solution entering
vat No. 3 was much reduced, and still more so the overflow into
No. 4, so that the solution seldom occupied more than three vats
besides the one which was undergoing charging. This gave
ample time to treat and discharge the residues, and to renew the
filter, if necessary, before the vat was again required in the cir-
cuit. By providing each tank with a sand filter only clear fil-
tered solution entered the precipitating vats, and no overflowing
solution at all.
When tank No. 1 was filled sufficiently with residues the
stream of pulp was changed to flow into vat No. 2 by opening the
corresponding plug-hole and closing the one above No. 1. The
communicating pipe between vat No. 1 and No. 2 was closed,
and the solution allowed to drain and then displaced by water.
The leaching with water was continued until the outflowing
stream did not show any reaction for silver. When a vat was
partly filled with residues the outlet under the filter could be
opened, giving a clear stream, though the vat was still under-
going the process of filling. When vat No. 1 was disconnected
from the circuit and No. 2 subjected to the operation of charg-
ing, the overflowing solution moved one tank farther and com-
menced to fill No. 5; when, in the course of the operation, the
overflowing solution reached vat No. 8, vat No. 1 was empty and
prepared to receive the overflowing solution. To prevent the
overflowing stream, when entering an empty tank, from washing
away the filter sand, a piece of filter cloth was spread over that
part of the filter and tacked to the inside of the vat, so that the
stream which flowed down on the side of the vat did not do any
harm to the filter.
Samples of the pulp taken at the drop showed that all the
silver chloride was extracted while the pulp was flowing through
the trough, and that the ore actually dropped as spent residues
into the vat. A similar experience was realized with regard to
the base-metal salts. After the pulp (ore and water) passed
TROUGH LIXIVIATION 243
through the short trough from the elevator to the sluicing-vat,
it was found by samples taken at the drop that all the heavy
metal salts soluble in water had been dissolved, and that only
some sodium sulphate still remained. Regular samples of the
base-metal solution were taken, but never found to contain any
silver.
TROUGH LIXIVIATION EXPERIMENTS ON A LARGE SCALE
While investigating the metallurgical problem of the lead- '
zinc ores of the San Francisco del Oro ore near Parral, Chihuahua,
Mexico, experiments were also made to treat the roasted ore by
the trough-lixiviation method. These experiments were made
previous to the working of the second-class ore dumps at Cusi-
huiriachic, described above. Experiences gained by this experi-
ment were used to advantage in working the "Cusi" ore.
The Bosque mill, near Parral, was very inconveniently
arranged. The main inconvenience was the want of grade;
therefore, the locality did not permit the erection of a complete
system for trough lixiviation, and the experiments had to be made
with only one circuit of six tanks, and I was obliged to use the
same troughs and tanks for base-metal and afterwards for silver
leaching. The washed ore had to be removed from the tanks
and brought to the head of the trough for silver leaching. Not-
withstanding this inconvenience, the experiments gave very
interesting results and information.
By a triangular trough, 138 ft. in length, f-in. fall per foot,
with a feed-box at the upper end, and intersected by five square
boxes, the pulp could be conveyed to any of the six tanks of the
circuit. The tanks were connected by pipes inserted near the rim.
The ore used in this experiment was roasted in the modified
Howell furnace. It was charged into a running stream of water
at the rate of 64 tons per twenty-four hours. The pulp passed
through the whole length of trough in fifty-five seconds.
In order to find out how much of the base-metal salts were
dissolved during this short time, and to ascertain the required
length of trough, samples were taken at different places, dried
and then subjected to a thorough washing in the laboratory,
with the following results:
Roasted ore before t roughing contained 12 per cent, in salts
soluble in water.
244 HYDROMETALLURGY OF SILVER
(1) The sample taken after the pulp passed the entire length
of 138 ft. still contained in salts soluble in water 4.9 per cent.
(2) The sample taken after the pulp passed through 58 ft.
of trough still contained in soluble salts 4.5 per cent.
(3) The sample taken after the pulp passed through 12 ft.
of trough still contained in soluble salts 3.6 per cent.
The above results are just in reverse order from what would
be expected; but it was not possible to take the sample from the
same portion of moving pulp, which may account for this irregu-
larity. Though this assumption is rather arbitrary, we may
accept it for want of a better explanation. If we take the aver-
age of the three results, we find that the pulp after troughing
still contained 4.7 per cent, of salts soluble in water, or, as the
roasted ore before washing contained 12 per cent, of such salts,
that 60.8 per cent, can be extracted while the pulp passes through
12 ft. of trough, or in 4.7 seconds. Long troughs are, therefore,
not essential for base-metal leaching. In order to ascertain if,
in tank Iixiviati6n in the usual routine, a larger percentage of
the soluble salts is extracted, a sample was taken from a tub
charge after it had been washed for eight hours and was
ready for silver leaching. The outflowing water gave with cal-
cium sulphide only faint white clouds, the usual indication that
base-metal leaching is completed. The sample, after drying and
weighing, was subjected to a second washing in the assay office,
and the result showed that of the original percentage of soluble
salts 61.7 per cent, were extracted by leaching in the tanks, which
is only 0.9 per cent, more than was extracted in the trough in 4.7
seconds.
In both cases about the same percentages of soluble salts are
retained by the ore, which only by a prolonged leaching can be
removed. They are not heavy metal salts, but principally
sodium sulphate and sodium chloride. In the present case
mostly sodium sulphate; for an analysis of the stock solution,
after three months' use, showed it to contain only 0.098 per cent,
chlorine, while the white clouds produced by an addition of cal-
cium sulphide proved to be gypsum.
TIME REQUIRED FOR BASE-METAL LEACHING
Though the dissolving of the base metals is almost instan-
taneous, considerable time is consumed in preparing the charge
TROUGH LIXIVIATION 245
for silver leaching, caused principally by the time required to
press out the base-metal solution by water; this time was found
to be 3 hours and 25 minutes for a charge of 8.39 tons. However,
the total time is still 3 hours and 35 minutes less than in tank
lixiviation.
The time is divided as follows:
IN TROUGH LIXIVIATION
hrs. min.
Leaching and filling the tanks 3 6
To drain remaining solution from top of the ore - 34
To press out base-metal solution by water 3 25
To press out water by hypo solution _! 20
Total time 8 h. 25 m.
IN TANK LIXIVIATION
hrs.
Charging 3
Base-metal leaching 8
Pressing out with hyposulphite solution 1
Total time 12 hrs.
QUANTITY OF WATER REQUIRED
Sufficient water had to be used to make the pulp move freely
through the trough, and to produce a sufficiently diluted base-
metal solution in order not to dissolve any silver chloride. The
results were obtained with 702 gallons per ton of ore, which is
equivalent to about one weight of ore to three of water. When
the tank was charged, and clean water turned on to press out
the solution, the speed of filtration was 12 inches per hour in a
tank of 10 ft 2 in. diameter, which is equivalent to 2065 gallons in
3 hours 25 minutes for a charge of 8.39 tons, or 246 gallons per
ton. After silver leaching, it took 2 hours 30 minutes to press
out the hyposulphite solution. Summing up, we find the total
consumption of water as follows :
gal.
In troughing 702
Pressing out the base-metal solution by water . 246
Pressing out the hyposulphite solution by water 181
Total consumption per ton 1129 = 150.5 cu. ft.
In tank lixiviation the consumption of water was found to be
703 gallons per ton of ore, or 93.7 cu. ft., which shows an increased
consumption in trough lixiviation of 56.8 cu. ft. per ton.
246 HYDROMETALLURGY OF SILVER
QUANTITY OF SILVER DISSOLVED BY THE BASE-METAL SOLUTION
One liter of the 702 gallons of base-metal solution was precipi-
tated with calcium sulphide. The precipitate, after fluxing and
treating like a common ore assay, returned not more than 0.0002
grams fine silver. If one liter contains 0.0002 grams silver, 702
gallons will contain 0.532 grams, which is the total amount of
silver dissolved from the whole charge of 8.39 tons of ore, or 0.06
grams, equal to 0.002 oz. silver per ton. This is practically
nothing, and the wash-water can therefore be allowed to run to
waste, without causing any perceptible loss in silver.
SILVER LEACHING
After three tanks were filled, and the base-metal solution
pressed out with water, and the water with a 0.38 per cent, sodium
hyposulphite solution, the ore was allowed to drain. Then the
ore was shoveled out and removed for silver leaching to the head
of the trough. Being at a time rather late in the evening, the
ore, saturated with hyposulphite solution, was left in a pile over
night. The next morning, however, it was found that some of
the silver chloride was decomposed during the night, and that the
chlorination test tailings had increased from 5.24 oz. to 9.03 oz.
per ton. Taking the tailings value of 9.03 oz. per ton as basis,
the experiment was continued. The measured stream of solu-
tion was kept uniform, while the rapidity of charging the ore
was changed according to the desired proportion. In order to
ascertain the proper length of trough, samples were taken at
different places, with the following results:
Assay office chlorination tailings of the pulp, 9.03 oz. per ton;
strength of solution, 0.38 per cent.; proportion, one weight of ore
to five of solution; rate of working, 38 tons of ore per day.
(1) Sample taken from spout of feed-box when entering the
trough: Tailings, 9.8 oz. per ton.
(2) Sample taken after passing 12 ft. of trough: Tailings, 8.13
oz. per ton.
(3) Sample taken after passing 70 ft. of trough: Tailings,
8.85 oz. per ton.
(4) Sample taken after passing 100 ft. of trough: Tailings,
8.13 oz. per ton.
TROUGH LIXIVIATION 247
(5) Sample taken after passing 120 ft. of trough: Tailings,
8.60 oz. per ton.
(6) Sample taken after passing 138 ft. of trough, while drop-
ping in tank: Tailings, 8.60 oz. per ton.
This experiment gave the very surprising information that
actually only a few feet of trough are required to produce a per-
fect dissolving of the silver chloride. It shows that the passing
through 12 ft. of trough, or in 4.7 seconds, the extraction is com-
plete, and that longer troughs are, therefore, not necessary.
This is of importance, as it simplifies the construction of trough-
lixiviating works and reduces the required grade.
OTHER PROPORTIONS, WORKING THE SAME LOT OF ORE
(7) Proportion; 1 ore to 3.4 solution; working rate, 55.8 tons
per twenty-four hours; tailings, 7.89 oz. per ton.
(8) Proportion, 1 ore to 2| solution; working rate, 84.5 tons
per twenty-four hours; tailings, 9.56 oz. per ton.
(9) Proportion, 1 ore to 10 solution; working rate, 19.05 tons
per twenty-four hours; tailings, 9.09 oz. per ton.
These results show that the proportion of 1 ore to 3.4 solution
gave the best results, the tailings being 1.14 oz. per ton poorer
than the chlorination assay called for.
A second series of experiments was made, and particular
attention was paid to avoid the decomposition of silver chloride.
The charges were subjected to silver leaching soon after being
saturated with hyposulphite solution.
Chlorination test tailings, 5.25 oz. per ton.
Strength of solution, 0.50 per cent.
(1) Proportion, 1 ore to 3.4 solution; tailings, 3.59 oz. per
ton.
(2) Proportion, 1 ore to 6 solution; tailings, 3.8 oz. per ton.
These are very satisfactory results; the tailings are as poor as,
in fact poorer than, those obtained in tank lixiviation after four
days' silver leaching. The proportion 1 to 3.4 proved again to
be sufficient, producing tailings 1.66 oz. poorer than the chlorina-
tion test called for; the quantity of solution required for this ore
is, therefore, very moderate, much less than that required in
tank lixiviation.
248 HYDROMETALLURGY OF SILVER
QUANTITY OF SOLUTION REQUIRED
By using the proportion 1 : 3.4 we need 100.8 cu. ft., or 816
gal. of solution to circulate for each ton of ore. In tank lixivia-
tion the required quantity for this ore was found to be 658 cu. ft.,
or 4935 gal. for each ton of ore, or about six times as much as in
trough lixiviation.
TIME REQUIRED FOR SILVER LEACHING
hrs. min.
Troughing and filling the tank 3 36
Draining the solution from the top of the ore - 34
Pressing out the solution with water ^2 30
Total time 6 h. 4m.
In the following we will compare the total time required by
the two methods from the time the ore enters the leaching works
until it is ready for discharge.
TIME REQUIRED IN TANK LIXIVIATION
hrs.
Charging 2
Base-metal leaching 8
Pressing out the water by solution 1
Silver leaching 96
Pressing out the solution by water li
Total time lOSYhours.
TIME REQUIRED IN TROUGH LIXIVIATION
hrs. min.
Base-metal leaching and filling the tank 3 6
To drain the wash-water from top of ore — 34
To press out the base-metal solution by water 3 25
To press out the water by hyposulphite solution ... 1 20
Silver leaching (sluicing with solution) 3 36
Draining solution from top of ore — 34
Pressing out the solution by water _2 30
Total time 15 h. 5m.
To work a charge of Del Oro ore by the trough system takes
15 hours 5 minutes, while by tank lixiviation it takes 108 hours
30 minutes, or about seven times as long.
FINENESS OF THE PRECIPITATE
The burned silver precipitate contained 20.9 per cent, of fine
silver, while that obtained in tank lixiviation during the same
week and from the same lot of ore contained only 17 per cent,
fine silver.
TROUGH LIXIVIATION 249
ADVANTAGES OF TROUGH LIXIVIATION
It can clearly be seen that trough lixiviation offers many
advantages over tank lixiviation, especially in large works, or if
badly filtering or slowly extracting ore has to be treated. In
large works, where twenty or more leaching- vats are in operation,
each one in. a different stage of the process, much attention is
required to avoid mistakes, while in trough lixiviation care has
to be given to only a few tanks. If the ore filters badly it will
take a very long time to extract the silver by leaching in tanks,
while in troughs the silver as well as the base-metal salts dissolve
almost instantaneously, and the effect of the bad filtering will
be felt only while the charge is draining and the solution is
being displaced by water. It makes the treatment of very badly
filtering ore possible, which otherwise could not be treated by
lixiviation.
The most important advantage of trough lixiviation is the
fact that this method enables the operator to bring the ore into
sudden contact with any desired quantity of the solvent. This
is a very important fact, as it offers the means to do away with
the special treatment of the base-metal solution, because a solu-
tion of this can be made sufficiently diluted not to dissolve
any silver chloride. Furthermore, the extraction of silver from
lead-bearing ores is slow, and requires an extensive plant. It is
well known to leachers of such ores that, while the main portion
of the silver is extracted in a short time, the remaining few ounces
will be tenaciously retained by the ore. Thus it happens that,
while the main portion of the silver can be extracted in the first
six or eight hours, the remaining eight or ten ounces of extract-
able silver will require three or four days, sometimes more. I
have seen cases in which only one ounce per ton in every twenty-
four hours of prolonged lixiviation could be extracted. This
very singular phenomenon is difficult to explain. It seems that
only that part of the silver is so difficult to extract which origi-
nally was contained in the lead-mineral of the ore. I have observed
that whenever the galena of the ore became richer in silver, or
had increased in quantity, the extraction became slow and drag-
ging. This, together with the fact that the main portion of the
silver can be quickly extracted, indicates that the slowness of
the extraction is not principally due to the disadvantageous in-
250 HYDROMETALLURGY OF SILVER
fluence of lead sulphate on the dissolving energy of the solution
for silver chloride, but that it must be due to some other cause,
which prevents the silver contained in the lead ore from dissolving
quickly except in large volumes of sodium hyposulphite solution.
In troughs, when the ore is brought at once into contact with the
required volume of solution, the silver dissolves almost instan-
taneously, and the presence of lead ore does not retard trough
lixiviation: it merely entails the use of larger quantities of solu-
tion, which is undoubtedly a very advantageous feature of this
method.
The cost of a plant is much less if arranged for troughs, as
large quantities of ore can be treated in a comparatively small
plant.
XVIII
THE RUSSELL AND KISS PROCESSES
(1) THE RUSSELL PROCESS
THE Russell process, which was *so elaborately and well written
up, and about which so many statements of excellent results
were published, and which was in consequence thereof introduced
at several places with a large expenditure of capital, has not
proved a success. As this process attracted much attention and
has found its way into all metallurgical text- and hand-books, it
is interesting and instructive to investigate the cause of the
failure.
It was claimed that the " extra solution " — a solution of a
double salt of cuprous hyposulphite and sodium hyposulphite,
manufactured by adding a solution of copper sulphate to a
solution of sodium hyposulphite — exerted a highly energetic
dissolving and decomposing action upon metallic silver, silver
sulphide, silver minerals belonging to the group of antimonial and
arsenical sulphides, and other silver combinations. Based on this
property of the " extra solution, " it was claimed that silver ores
treated by this process required a less careful chloridizing roast-
ing, or only an oxidizing roasting, and that even raw sulphureted
ores could be successfully desilverized.
These claims attracted general attention, and if they had
been true in the sense in which they were given out this process
would have marked a decided step forward in the hydrometallurgy
of silver; but these claims were based on results obtained on a
very small scale, and under conditions which are not practicable
to create and maintain on a large scale, while the results, even
under such conditions, especially with regard to raw ores, proved
to be not good enough to justify the application of the process in
practical metallurgy. To protect himself against failure it is of
the greatest importance that the experimenter should execute
251
252 HYDROMETALLURGY OF SILVER
his experiments under conditions which can be practically main-
tained on a large scale. As soon as he has, by working on a large
scale, to change the conditions wrhich he maintained in his labora-
tory tests, he will obtain very different results, and only too
often will experience a complete failure.
If a half ounce of raw sulphureted ore is treated in a beaker
with a large excess of a 32 per cent, "extra solution" for twelve
hours with frequent stirring, as was done by Russell in the labora-
tory, it is experiment ing under condit ions which cannot be profitably
maintained on a large scale, and the results thus obtained, if pub-
lished, should be given as a matter of scientific interest but not as
actual results of a new metallurgical process, especially if the con-
dition under which these results were obtained are withheld. More
or less silver will be dissolved by such a strong "extra solution";
a 32 per cent, sodium hyposulphite solution applied in the same
manner will also dissolve some silver; but it is not practicable to
work with such concentrated solutions, since they cannot be main-
tained without a very large consumption of sodium hyposulphite
and copper sulphate. Besides, raw sulphureted ores do not filter
well, and the ore would have to be treated in agitating tanks for
twelve hours, and the separation of the solution from the residues
would have to be effected by means of large filter-presses. These
manipulations, together with the very large consumption of
chemicals, would cost more than roasting with salt. The weakest
part of this method as applied to raw ore, however, is that the
extraction at the best is so inferior and incomplete that its appli-
cation is entirely out of the question.
More favorable results can be obtained by treating raw oxi-
dized ores, or sulphureted ores which were first subjected to a
thorough oxidizing roasting, wherein there are not so serious dif-
ficulties preventing its application on a large scale, though only
in exceptional cases will it be rational to employ it. A dilute so-
lution can be used, and often 50 to 70 per cent, of the silver will
be extracted. But similar results can be obtained by using a
straight solution of sodium hyposulphite. Some of the San
Francisco del Oro ore, a highly sulphureted lead-zinc ore, was
roasted oxidizingly in a reverberatory furnace by me, and part
of it treated with sodium hyposulphite and part with Russell's
"extra solution." The roasted ore contained 29.3 oz. silver per
ton. By leaching with sodium hyposulphite 17.2 oz. silver were
THE RUSSELL AND KISS PROCESSES 253
extracted, while with Russell's "extra solution" the extraction
gave 17.64 oz. or 0.43 oz. silver per ton more. But an increased
extraction of less than half an ounce of silver per ton does not
justify the extra consumption of 7 Ib. of copper sulphate and 5
Ib. of sodium hyposulphite, which amount is stated to be the
consumption of chemicals in the Russell process per ton of ore
treated. It would be folly to oxidize an ore for lixiviation and
obtain an extraction of only 50 to 70 per cent, instead of chlori-
dizing it and obtaining an extraction of 90 per cent, and more,
especially since by the modern method of chloridizing roasting
the loss of silver by volatilization is greatly reduced, and does
not exceed the loss occurring in oxidizing roasting. The claim
that ores require only an oxidizing roasting, if treated by the
Russell process, is therefore a mistake, and cannot be verified by
actual and satisfactory working results.
It was mentioned above that only in exceptional cases can
lixiviation of raw ores be executed successfully. One of the
cases is when in oxidized ore the silver occurs as chloride. How-
ever, it is more advantageous to subject such ore first to a short
red heat to melt the silver chloride. Chloride of silver, as it
occurs in nature, is very dense and dissolves very slowly, but if
the ore is heated the silver chloride will melt and impregnate the
surrounding ore particles, in which condition it offers a large sur-
face to the solvent and permits a much quicker extraction. Oxi-
dized ores in which the silver occurs as antimonate can also be
successfully leached. Low-grade oxidized ores, which will yield
60 to 70 per cent, by leaching raw, can also be successfully treated
if, on account of the small tenor of silver in the ore, the increased
amount extracted by chloridizing does not exceed the cost of
roasting. However, the use of sodium hyposulphite for such
ores will be found more economical than the use of the "extra
solution," the additional cost of chemicals not being covered by
the slight gain in extraction.
After the process was tried on a large scale without success,
Russell modified his process. He adopted chloridizing roasting
and leaching with straight sodium hyposulphite, and after this
was done applied his "extra solution," by which he claimed to
obtain a much better extraction. But this assertion was again
based on laboratory experiments, which were made under dif-
ferent conditions than could be maintained in the works. Based
254 HYDROMETALLURGY OF SILVER
on these results it was next claimed that silver ores require a
less careful chloridizing roasting, because the "extra solution"
dissolves the unchloridized part of the silver. But this claim
was also found to be an illusion, as has been demonstrated by
many failures on a large scale. The following case is an illustra-
tion:
In Sombrerete, Zacatecas, Mexico, the lixiviation process
with sodium hyposulphite was in successful operation for years,
until a new company was organized to work the property on a
larger scale. A new and large mill was erected to suit the re-
quirements of the Russell process. The success in the old mill
was based on a good chloridizing roasting in reverberatory fur-
naces. The results, however, were entirely different when the
new mill was set in operation. In this the ore was roasted in
a Stetefeldt furnace, which is not suitable for such heavy sul-
phureted ore, and the chlorination was far from being satisfac-
tory. This gave an opportunity to demonstrate the claim that
the "extra solution" exerts an energetic dissolving and decom-
posing action on the unchloridized part of the silver. However,
the "extra solution" failed to react, and after eleven months of
unsuccessful trials the company failed and the property changed
hands.
The new company abandoned the Stetefeldt furnace and the
Russell process, built a suitable number of reverberatory fur-
naces, and adopted the common lixiviation process with sodium
hyposulphite with great success. The mill is still in operation
and treats regularly 60 to 80 tons of ore per day.
The Cusi company at Cusihuiriachic, Chihuahua, Mexico,
had a similar experience. The old lixiviation process was for
years in successful operation on a scale of 50 to 60 tons per day.
Induced by the glowing representations of the advantages of
the Russell process, the company adopted it, but after 1J years'
trial, with heavy financial loss, the process was discarded and
lixiviation with sodium hyposulphite was resumed.
(2) THE Kiss PROCESS
This process was recommended for the mutual extraction of
silver and gold from sulphureted auriferous silver ores. Kiss
subjected the ore to a chloridizing roasting, leached with water
to remove the base-metal chlorides, and extracted the silver and
THE RUSSELL AND KISS PROCESSES 255
gold with a solution of calcium hyposulphite. As precipitant
he used calcium polysulphide. According to Kiss, in chloridizing
roasting subchloride of gold is formed, which is neither soluble in
water nor in a solution of sodium chloride, but is soluble in a solu-
tion of calcium hyposulphite. This method was in actual opera-
tion in several places in Hungary, and while the extraction of
the silver was satisfactory (90 per cent.) the extraction of the
gold varied greatly at different places; from 90 to 20 per cent.
This was undoubtedly caused by the way the roasting was done
at the different places. Subchloride of gold cannot exist at a
temperature prevailing in the furnace, especially not if the tem-
perature toward the end is increased, as was formerly generally
done; it is then decomposed into metallic gold and chlorine.
The subchloride of gold, however, is formed if the ore after being
discharged from the furnace is allowed to cool slowly. If the
hot ore on leaving the furnace is dumped into a pile, or better in
a bin, the generation of chlorine continues down to a tempera-
ture which is low enough not to decompose the subchloride of
gold, and therefore opportunity is given for the formation of
this gold combination. (This has been discussed in another
chapter of this treatise.) The variation of the gold extraction
at the different places was therefore most likely due more to
whether the chloridized ore was cooled slowly or quickly rather
than to the different character of the ores.
We have seen, too, that calcium hyposulphite in contact with
sodium sulphate forms sodium hyposulphite and calcium sul-
phate (gypsum), which precipitates, so that Kiss actually did not
work with calcium hyposulphite after the solution was used for
some time, but with sodium hyposulphite, as the lime salt was
decomposed by coming in contact with the remaining sodium
sulphate of the roasted ore. It is questionable whether calcium
hyposulphite is a more energetic solvent for subchloride of gold
than sodium hyposulphite; at least, it was not proved by Kiss'
method, as he actually leached with the soda and not with the
lime salt.
XIX
THE AUGUSTIN PROCESS
IN this process the material is roasted with salt. The
resulting silver chloride is dissolved by a hot concentrated brine.
From the solution the silver is precipitated by copper, and the
copper by iron. The remaining solution is freed from iron and
sodium sulphate and brought into circulation again.
The ore is roasted, first oxidizingly to transform the metal sul-
phides into oxides, and the silver sulphide as much as possible
into silver sulphate, so that during the chloridizing period but
little volatile or soluble metal chlorides will be formed. The
material which conforms best with these requirements is copper
matte. The oxidizing roasting is continued until the iron has
changed into red oxide and the copper into cupric oxide. If no,
or not enough, sulphates are left at the end of the oxidizing period
to decompose the salt, ferrous sulphate has to be added
with the salt to generate the chlorine. The roasted material is
not first leached with water, but the hot concentrated salt solu-
tion is applied at once to the ore. If not all the copper was
changed into oxide, and some chloride or subchloride is still pres-
ent, this will be dissolved by the salt solution, and the subchlo-
ride oxidizes during the precipitation of the silver to basic insoluble
chloride of copper and makes the precipitated silver impure.
If the matte contains lead the hot brine will dissolve lead chloride,
which in cooling precipitates as white flakes and also makes the
silver impure. Such material has to be leached first with hot
water.
The stream of hot brine coming from the leaching-vats is
divided in many small streams, each of which is made to strike
metallic copper, which is placed in small wooden tubs provided
with a filter and an outlet below the filter. Under these tubs is
placed another row of similar tubs, and below this still another,
256
THE AUGUSTIN PROCESS 257
so that the brine passes through quite a number of such tubs
filled with copper. They are arranged on benches. Leaving
the last row, the desilverized solution is conveyed to a series of
tanks for the precipitation of the copper by scrap iron.
An advantageous feature of this process is that the silver is
precipitated in the metallic state; from a solution of sodium hypo-
sulphite, the silver could not be precipitated with copper without
a decomposition of the sodium hyposulphite. This method,
however, has many disadvantages which make it unfit for work
on a really large scale. The dissolving energy for silver chloride
of a concentrated hot salt solution is much inferior to that of
sodium hyposulphite; it takes 68 parts of sodium chloride to dis-
solve one part of silver chloride, while it takes only two parts of
sodium hyposulphite. The handling of large quantities of hot
concentrated brine is rather unclean and troublesome. All parts
of the vats and tubs above the solutions and the whole of the
outside become covered with incrustations of salt; besides, it is
exceedingly difficult to keep them tight, as such a concentrated
and hot solution finds its way even through the fibers of the wood.
Cement copper acts more energetically in the precipitation of
the silver than scrap copper, but the resulting cement silver is
not quite as clean, as it usually contains more copper particles,
which have to be removed by treating with hydrochloric acid.
XX
EXTRACTION WITH SULPHURIC ACID
THIS process, like the Augustin method, is exclusively used
for the extraction of the silver from products of smelting, prin-
cipally black copper and copper matte.
(1) EXTRACTION OF SILVER FROM COPPER MATTE
The products of this process are cupric sulphate (blue vitriol,
blue stone), which goes in solution, and residues in which the
silver and gold remain, from which they are finally extracted by
smelting with lead ores.
(a) The Old Method
Besides the extraction of the precious metals, it is of great
importance that the resulting cupric sulphate should be as free
from iron as possible, because the blue vitriol is bought in the
market as such, and if it contains ferrous sulphate its value is
much reduced. To accomplish this, the common copper matte
has to be first concentrated by repeated roasting and smelting,
not only to enrich it in copper but to free it as much as possible
from iron. Though red oxide of iron dissolves much more
slowly in diluted sulphuric acid than cupric oxide, the resulting
copper solution will contain by far too much ferrous sulphate to
produce a marketable blue vitriol if the copper matte contains
a large percentage of iron.
At Freiberg, Saxony, where this process is in operation, they
concentrate the original copper matte of 38 to 44 percent, copper
by repeated roasting and smelting until it contains 70 or more
per cent, copper and only 0.3 per cent. iron. This matte is crushed
coarse (J-in. mesh) and roasted twelve hours in a reverberatory
furnace. The charge of 1000 Ib. is continuously stirred, then
258
EXTRACTION WITH SULPHURIC ACID 259
withdrawn from the furnace, finely pulverized and subjected
during four hours to a "dead roasting." The roasted matte is
sifted, the coarse pulverized and mixed with the fine.
For the treatment of the roasted matte with diluted sulphuric
acid there are cylindrical upright drums 2 ft. 8 in. in diameter
and 3J ft. high, made of i-in. lead supported by an iron frame-
work, whch is covered with thin sheet lead. The drum has two
outlets, a lower one for discharging the residues and one higher
up for the discharge of the copper solution. Extending to near the
bottom there is a lead pipe for the injection of superheated
steam. This pipe is so arranged that it can be lowered and
raised. The drum is charged with 3 cu. ft. of mother liquor,
and 3 cu. ft. of common sulphuric acid of 45 to 47 deg. B.,
which gives a solution 34 to 36 deg. Beaume". Superheated steam
is now applied, and when the solution is boiling, 200 Ib. of roasted
matte is gradually charged with a hand shovel through a movable
funnel. The pulp has to be charged continuously. After a
quarter of an hour, 9 cu. ft. of mother liquor is added to the
contents of the drum, which fills it to about 7 in. below the rim.
The boiling with steam is continued for 4 to 5 hours, during which
time stirring has to be done at intervals. The steam is turned off,
and half an hour is given for the residues to settle. Then the
solution is drawn off through the upper outlet and conveyed to
settling tubs, while the residue, containing all the silver and
about 5 per cent, copper, is transferred into a special basin, where
it is washed with water. The still hot copper solution is left
about an hour in the settling-tanks, during which time some basic
iron salts settle. Then it is drawn off into the crystallizing tanks.
The residues are dried and transferred to the lead smelting.
The blue vitriol crystals after five to seven days are removed
from the crystallizing tanks. They are not of a clear blue color,
but have a greenish appearance from a certain percentage -of iron
which they contain, and are again dissolved and recrystallized.
All the solutions are very acid.
The yearly production of blue vitriol resulting from this process
is about 800 tons.
(b) 0. Hofmann's Method
We have seen that, to treat copper matte with sulphuric
acid by the old method, it is required to first eliminate nearly
260 HYDROMETALLURGY OF SILVER
all the iron from it by repeated roasting and smelting, which at
the same time produces a matte very rich in copper. This
careful elimination of the iron is done to permit the production
of a purer blue vitriol. We have also seen that in order to avoid
a concentration of the resulting cupric sulphate solution for
crystallization the proportion of acid and roasted matte is so regu-
lated as to produce at once a solution strong enough for that
purpose. If this is done for economical reasons, to avoid the
cost of evaporation, the object is not attained, as it necessitates
recrystallization because the crystals contain too much iron, not-
withstanding the fact that the matte by previous treatment con-
tained but very little iron. The solubility of red oxide of iron in
sulphuric acid increases with the strength of the acid, therefore
it is not advisable to bring the matte in sudden contact with too
strong an acid. Cupric oxide dissolves so much easier than the
iron oxide that a weak acid can be used which will readily dis-
solve the cupric oxide, but very little of the iron oxide, and the
resulting cupric sulphate solution will be much purer. The
strong first solution makes this method not suitable to be worked
on a really large scale, because concentrated solutions require
but little cooling to form fine crystals, which are very annoying in
the separating of the residues from the solution, because these
crystals will form in the filter-press, clog the filter cloth and
prevent a free filtration. They will form in the mass of the
residues, whence they can be removed only by a prolonged wash-
ing with hot water, causing the formation and accumulation of
too large quantities of a very weak solution.
I discovered a method of purifying a cupric sulphate
solution from iron, antimony, arsenic, etc., which, together with
a modification of the manner in which the process of dissolving
the cupric oxide is executed, made the sulphuric acid process not
only much cheaper, but also applicable for working common
copper matte without previous concentration, at the same time
producing a very pure product of blue vitriol.
The method is based on the reaction that, if through a hot
neutral solution of cupric sulphate which contains ferrous sulphate
a stream of air is forced, and at the same time cupric oxide is added,
ferric oxide is precipitated and cupric sulphate will go in solution.
Instead of specially prepared cupric oxide, roasted copper matte
is used, which contains sufficient cupric oxide for the reaction.
EXTRACTION WITH SULPHURIC ACID 261
The manner in which the dissolving of the roasted matte is
performed is modified, inasmuch as the roasted matte is not added
to the bulk of acid of sufficient strength to produce at once a
cupric sulphate solution of the required concentration, but the
solution is gradually enriched in copper until the desired degree
is reached. This is done by preparing first a 3 per cent, acid
solution. Such a weak acid dissolves but very little iron oxide.
Then, when hot, the copper is added gradually as a stream of
roasted matte, while at the same time a small stream of 60 deg.
B. acid flows into the tank. The solution in the tank is agitated
vigorously by a machine-stirrer. The proportion of matte and
acid has to be so regulated that the solution always maintains
its strength of 3 per cent, in acid, while the content of copper
increases. This is controlled by frequent volumetric tests.
When the desired strength in copper is nearly attained, the influx
of acid is stopped and only so much matte added as will neutralize
the 3 per cent. acid. Thus the copper matte does not come in
contact with a stronger acid, and the resulting solution does not
contain more than 0.7 to 1 per cent, iron, notwithstanding that
the treated material is very rich in iron oxide.
This process was introduced by me in the large smelting
works of the Consolidated Kansas City Smelting and Refining
Company, at Argentine, Kansas, for the treatment of the
argentiferous leady copper matte, first on a medjum large scale,
but on account of its successful working it underwent an enlarge-
ment every year, until a daily producing capacity of 60 tons of
blue vitriol was reached. In the following a short description
of the process and the apparatus will be given.1
The material treated is a leady copper matte containing 34 to
40 per cent, copper and 12 to 14 per cent. lead. It is crushed first
in a rock breaker and then pulverized in a Krupp ball-mill of 100
tons daily capacity through a screen of 50 meshes to the lineal inch.
The roasting is done in three two-story Pearce turret furnaces,
each provided with three fireplaces at the lower and two at the
upper hearth. For the benefit of the subsequent operation the
roasting has to be conducted with great care, and with attention
1 A more elaborate description can be found in " The Mineral Industry,"
Vol. 'VIII and Vol. X. The plant is not now in operation, the smelting
works with which it was connected having been closed down and dismantled
in 1902.
262 HYDROMETALLURGY OF SILVER
equally divided between* the oxidation of the copper and of the
iron. The copper is to be converted into cupric oxide and sul-
phate, and the iron into red oxide, in which state it dissolves only
slowly in hot diluted sulphuric acid. The formation of cupric
sulphate is very desirable, as it saves acid in the subsequent
treatment, but still it is not advisable to conduct the roasting so
that as much as possible sulphate is formed, because the roasted
material will then contain too much soluble iron, which would
make the resulting cupric sulphate solution too impure. At a
comparatively early stage of the roasting nearly all the copper
is in a state in which it can be extracted by diluted sulphuric
acid; about 75 per cent, of it is present as oxide and 25 per cent, as
sulphate. The roasting, however, cannot be considered complete
at this stage, because the roasted material contains yet too much
soluble iron.
At the beginning of the operation the temperature should not
be raised above that produced by the combustion of the sulphur
of the matte. During this period the charge assumes a rather
bright red appearance, an effect due more to light than to heat,
and if excess of air is admitted to the furnace but few lumps
will form, even with very leady copper matte. Gradually, as
the oxidation advances, the surface of the charge becomes darker,
especially near the air-doors, and when the entire surface of the
charge begins to darken, the fire is slightly increased to prevent
cooling, as from this time on the supply of heat furnished by the
oxidation decreases rapidly. If this condition is overlooked
and the charge cools too much, it will be difficult to raise the
temperature again to the proper degree.
The roasting is continued at a moderate temperature until
no more heat is evolved by the oxidation of the material, after
which the temperature must be raised to a cherry red. A skilled
roaster can readily determine this point by stirring the charge;
if the particles thrown to the surface brighten, the roasting has not
advanced far enough, but if the entire charge presents a dead and
uniform red color, it shows that this part of the roasting has been
completed, and that it is time for an increase of temperature.
This can now be done without danger of lumping the charge,
because by this time the greater part of the sulphides has been
oxidized. The increase of the temperature is necessary for two
reasons : first, to hasten the oxidation of the remaining sulphides,
EXTRACTION WITH SULPHURIC ACID 263
which would require a very long time at a low temperature, and
second, in order to decompose the iron salts and to convert them
into the red oxide. The task for the roaster is to convert as much
iron as possible into the red oxide without decomposing the
cupric sulphate present. Cupric sulphate resists considerable
heat, more so than the ferrous salts, and it is possible to conduct
the roasting in this way; but the increase of temperature requires
judicious care, because if the heat is too high the cupric sulphate
will be reduced to cuprous oxide, in which condition but half of
the copper is soluble in diluted sulphuric acid. If crystals of
cupric sulphate are exposed to heat and air, it will be noticed
that after the acid is expelled the mass assumes a red color
showing the formation of cuprous oxide. If heating is continued,
it turns black by being oxidized to cupric oxide. Should cuprous
oxide be formed, the amount of extractable copper will be greatly
reduced. If, for instance, the extractable copper before raising
the temperature was 97 per cent., after an excessive heat it will
be reduced to 92 or even 90 per cent. When the roasting is
done in a common reverberatory furnace, a mistake of this kind
can be corrected by keeping the charge longer in the furnace and
thus oxidizing the cuprous to cupric oxide. In a mechanical con-
tinuously discharging furnace, like the Pearce, however, this can-
not be done, but with experience and care the decomposition of
the cupric sulphate can be avoided. It is of great importance
that, during the whole time of roasting, air has free access into
the furnace.
It is not possible to avoid the formation of some lumps, espe-
cially in roasting leady matte, but if the roasting is conducted
properly, these will be small, soft, porous, and consist of well
roasted material. Roasted matte is always of coarser grain than
the raw pulp, and for this reason as well as on account of the
lumps it is necessary to pulverize the roasted material before
treatment with sulphuric acid. This is best done in a Krupp
ball-mill through a screen with 50 meshes to the lineal inch.
The dissolving at Argentine, Kansas, is done in eight agitating
or stir tanks, which are arranged in two parallel rows with a
track between for the delivery of the matte. The cars are scoop-
shaped and are partly covered with sheet iron, and so made that
the cover and scoop end come close together, leaving only a nar-
row slit open, so that when the car is tilted the roasted matte
264
HYDROMETALLURGY OF SILVER
runs gradually into the dissolving stir tank. Fig. 66 represents
the vertical section of a stir tank, 12 ft. in diameter and 6 ft. deep.
The bottom and sides are lined with boards for protection against
wear from the friction of the pulp. In the center is suspended a
Hard Maple
/* Be
Bearing Blocks
6ilOy.p.
FIG. 66. — STIR TANK, VERTICAL SECTION.
heavy wooden shaft, having fastened to the lower end, by heavy
brass plates, a cross-beam cut at both ends like a propeller-blade.
In the center, below the propeller and fastened to the bottom of
EXTRACTION WITH SULPHURIC ACID 265
the tank, is a cone-shaped projection constructed of strong wooden
staves, the interior of which is filled with sand. This cone forces
the matte, when the propeller is in operation, toward the pe-
riphery, where it is subjected to the swift rotating motion of the
liquor, thus preventing its accumulation in the central part,
where the motion is much less. Two outlet pipes provided with
hard-lead valves make connection with the pressure tank. These
are placed one above the other to permit the withdrawal of a'
portion of the clear liquor if desired, in which case the paddle is
stopped and the residues allowed to settle. In the usual working
of the tank, however, the lower outlet pipe only is used. The
wooden ring attached to the rim of the tank prevents the splash-
ing of the pulp. The tank and trestle support are placed in a
flat lead pan to collect any leakage. The upper part of the shaft
is provided with a gear-wheel to receive the power.
The stir tanks are filled about two-thirds full with water, the
agitator set in motion, and sulphuric acid added until the liquid
shows about 3 per cent. acid. The matte is then charged gradually,
and at the same time a stream of acid is allowed to flow in so as
to maintain the same acid strength during charging. In this
way the dissolving is accomplished with an acid strength of 3
per cent, or less, and still yields a strong solution of cupric sul-
phate. As stated above, it is preferable to work with weak acid,
because much less iron and other impurities will be dissolved than
with stronger acid. When the solution has attained a strength
of about 20 to 22 deg. B., the flow of sulphuric acid is stopped
and matte only charged until the solution is neutral. Toward
the end it is advisable to charge the matte at intervals, and to
make frequent acid tests to avoid an excess of matte. The addi-
tion of sulphuric acid to the water, and the matte to the diluted
acid, produces heat, which aids the solution of the cupric oxide.
This heat is not sufficient, and the temperature is further raised
by a jet of steam. It is well to interrupt the charging of matte
while neutralizing as soon as the solution shows 1 per cent, free
acid. If the agitation is then continued this 1 per cent, of acid
will be diminished, but if in half an hour after the last acid test
the percentage of free acid remains the same, more matte is
added until the solution is neutral; in this way a mistake of add-
ing a large excess of matte is avoided.
Below the stir tanks are the pressure tanks, into which the
266
HYDROMETALLURGY OF SILVER
finished charge is drawn while the paddle is in motion. On
account of the residues the pressure tanks are in an upright
position, and are constructed as follows: The body consists of
two cylindrical sections 4 ft. long each and 4 ft. 6 in. in diameter,
the bottom being of a spherical shape of 2 ft. 3 in. radius. The top
is rounding upward to a hight of 6 in. above the rim. The sections
are tightly flanged, with a rubber gasket between. Four pipe
connections are made through the top; the discharge-pipe enter-
ing through the center passes nearly to the bottom, and the filling-
FIGS. 67, 68, 69. -CAST-IRON PRESSURE TANK.
67 (top right), top view. 68, section on line AB. 69, plan of supporting frame.
pipe, air-inlet and air-outlet pipes are conveniently arranged
around it. By proper connection with the filling-pipe and by
the use of valves, one pressure tank is made to serve four stir
tanks. The upper cylindrical section is provided with a manhole,
and the tanks are made of cast iron and lined first with lead and
then with wood to protect the lead from wear by abrasion. The
air pressure required is from 40 to 50 Ib. Fig. 67 shows the top
view and Fig. 68 a vertical section of an upright pressure tank
EXTRACTION WITH SULPHURIC ACID 267
while Fig. 69 illustrates the supporting frame. This tank as
represented by the figures is smaller than those used at Argentine,
and has only one cylindrical section instead of two. The general
arrangement, however, is the same. The air-inlet extends down-
ward along the side, ending near the discharge-pipe, in order to
keep the residues near the outlet in a loose condition and prevent
the pipe from clogging. For this reason it is advisable to have a
small stream of air enter during filling.
The pulp is forced by the pressure tank into a large filter-
press, of which there are five in use. The press is 25 ft. long and
has hardwood frames and plates, and holds, when filled, 5 tons of
residues. The clarified solution flows from the press to lead-
lined storage tanks, from which it is elevated, by means of a
pressure tank, to the purifying towers. These towers were origi-
nally made of 20-lb. sheet lead incased in a cast-iron framework,
which, however, were replaced by larger ones made of 4-in. Cali-
fornia redwood. Fig. 70 represents a vertical section of such a
tower. They are made of staves 16 ft. long, 9 ft. in diameter, and
are well hooped with round iron. The top and bottom are flat.
The towers are firmly fastened to a strong wooden trestle, which
in turn is anchored to a concrete foundation. This construction
is called for to guard against the oscillating movement of the tank
from the action of the contained solution, which is set in violent
motion by the ascending air. Inside the tower, about 19 in.
above the bottom, the 4-in. lead air pipe enters, and is connected
with a perforated 6-in. lead pipe, which extends diametrically to
the opposite side and is closed at the further end. Outside the
tower the pipe extends upward to above the top of the tower,
and thence downward to the discharge-pipe of an air-compressor
or to an air-receiver, connected with the compressor. This
arrangement is necessary to prevent the solution from flowing
to the air-pumps if the latter be not in operation. At the same
level with the air-inlet, the 4-in. discharge-pipe enters, which is
provided with a hard-lead valve placed close to the outside of
the tower. At about the same level enters a 1-in. lead steam-pipe
for heating the charge. The direct application of steam does
not dilute the solution, because the evaporation, which is favored
by the ascending air, fully equalizes the dilution by condensed
steam. The top of the tower is provided with an 8-in. pipe ex-
tending through the roof for the escape of steam and air, which,
268
HYDROMETALLURGY OF SILVER
FIG. 70.— TOWER FOR REFINING CUPRIC SULPHATE SOLUTIONS.
EXTRACTION WITH SULPHURIC ACID 269
however, first are made to pass through a lead-lined box provided
with shelves arranged in a zigzag manner to precipitate all par-
ticles of liquor which may be carried out by the current of air.
A 4-in. pipe also enters the top, which serves for filling the tower
with crude copper solution. An opening in the center is connected
with a small hopper which is filled with roasted matte. The upper
third part of the tower is provided with glass tube gages to con-
trol the filling. The perforations of the air-pipe are only on the
bottom side, to prevent the inflow of matte and precipitate. The
steam pipe enters through the lid of the manhole, as the heat of
the pipe affects the wood at its immediate surrounding.
The tower is filled with crude solution, leaving, however, suf-
ficient room for the increase in volume of the solution, which
immediately takes place as soon as the compressed air is supplied.
A tower can be charged with about 5000 gallons of solution, and
as three such charges can be refined in twenty-four hours the
working capacity of one tower is 15,000 gallons per day. At
Argentine eight such towers are in use. When the tower is filled,
steam is allowed to enter, and also some air, to produce a more
uniform heating of the solution. The air causes the precipitation
of some basic iron salts, but I never succeeded in precipitating
more than about half of the iron the solution contained, although
the treatment was extended many hours. When the solution
is hot (75 to 80 deg. C.) more air is admitted, and some roasted
matte from the hopper is made to drop into the tower. The
violent boiling motion of the solution keeps the matte in suspen-
sion, and after three to four hours the solution will be entirely
free from iron, arsenic, antimony, etc. To observe and to regu-
late the progress of the operation the solution is tested from time
to time for iron by taking samples through a small cock inserted
in the side of the tower. If between two tests the content of
iron is not diminished some more matte is added. It is not neces-
sary to test for other impurities, because the iron predominates,
and by the time all of it has been precipitated, no trace of any
other will be found.
The cupric oxide in presence of air combines with the sul-
phuric acid of the ferrous sulphate, forming cupric sulphate,
while the iron precipitates as oxide — a decided advantage, as
the precipitant is converted into cupric sulphate, and thus
enriches the solution in copper.
270 HYDROMETALLURGY OF SILVER
The refined solution leaving the tower should not be stronger
than 24 to 26 deg. B. when hot, as otherwise it will cause trouble
in the filter-press, for reasons above explained. Should a con-
centration have taken place in the tower beyond this, the solution
should be diluted. The tower is discharged into a special stir
tank and from there, by means of a pressure tank, forced into a
filter-press. The clear and purified liquor is conveyed to the
evaporating department, while the residues are subjected to an
additional treatment. These tower residues are of a grayish
yellow color and consist principally of precipitated iron, arsenic,
antimony and some undecomposed matte, with also some basic
copper sulphate. To remove the last named substance the resi-
dues are treated in a1 stir tank with 2.5 to 3 per cent, cold acid
solution, which dissolves the basic copper salt, leaving the im-
purities unaffected, with the exception of iron, which is acted on
very slightly.
The refined solution does not contain even a trace of silver.
Whatever silver is converted into sulphate during the roasting
is precipitated in the stir tanks by the ferrous sulphate present
in the solution. The residues are washed well in the filter-press
to remove all copper solution. The wash-water is collected
separately, and used instead of fresh water for the preparation
of a new charge in the stir tanks. The residues contain all the
silver, gold and lead of the roasted matte; they are of a dark-red
color and consist mainly of iron oxide. They are sent to the
lead-smelting department for the extraction of the precious
metals and the lead. On account of the large percentage of iron
oxide and lead the residues are an excellent material for lead
smelting. A large portion of the lead of the matte is converted
into sulphate by the roasting, and therefore does not act much as a
consumer of acid. As the roasted matte contains 20 to 25 per
cent, of its copper as sulphate, the total consumption of acid is
much less than the equivalent amount contained in the blue
vitriol produced.
The evaporating department furnishes 90,000 gallons of
concentrated solution daily. To supply this amount an im-
provement over the old pan evaporator with under-fire or steam
coils was requisite, and as vacuum evaporators could not be
adopted I finally devised and introduced an economical and
effective evaporator, which is illustrated in Figs. 71 and 72. The
EXTRACTION WITH SULPHURIC ACID
271
272 HYDROMETALLURGY OF SILVER
principle observed in the construction of this evaporator is the
application of the hot furnace gases in a manner by which almost
a complete utilization of the heat contained in them takes place.
The apparatus consists of a flat tank, wooden with the exception
of a 2-ft. space at the front end, which is made of steel, so that the
wood will not be in too close proximity to the furnace. The
tank is 65 ft. long, 12 ft. wide and 2 ft. deep, and is lead lined,
the two ends having much heavier lead lining than the sides and
bottom. It is traversed longitudinally by 13 6-in. heavy lead
pipes. These pipes rest on bricks which are properly placed on
the bottom of the tank. The tank rests on wooden trestle-work
of a proper hight to correspond with the hight of the furnace.
At the furnace end in each lead pipe is inserted a 5-in. pipe 4 ft.
long, provided at the outer end with a flange. The iron pipe
serves to protect the lead pipe from immediate contact with the
red-hot gases from the furnace. They also make the connections
between the lead pipes of the tank and the iron pipes of the
furnace.
The furnace is comprised of the fireplace C, the dust-chamber
D and the distributing chamber D1. At a proper hight in the
wall of the distributing chamber nearest the tank are 13 open-
ings, in each of which is inserted a short cast-iron pipe, 5 in. in
diameter, with a flange at the outer end. Each pipe is connected
with its corresponding piece of iron pipe inserted in the lead pipe
of the tank by a cast-iron S-shaped elbow, Y, which allows the
introduction of the compressed-air pipe E for removing any
accumulation of ashes in the lead pipes of the tank.
An American "underfed stoker" is used for the slack-coal fuel
and affords practically perfect combustion, which is of great
importance, as otherwise the lead pipes would soon become
coated with soot and lose much of their efficiency to transmit the
heat to the solution.
The opposite ends of the lead pipes in the tank are connected
with the brick suction chamber 0, which in turn is connected by
a galvanized iron pipe, R, with a suction fan, P, the gases being
discharged into an underground flue, S. This flue serves in com-
mon to collect the waste gases from 11 evaporators, and terminates
outside the building in a brick chimney 40 ft. in hight.
The top of the pan is closed with a wooden cover, and wooden
joists, G, are placed across the pan about 5 ft. apart, having
EXTRACTION WITH SULPHURIC ACID 273
cleats fastened to the lower side, as shown in Fig. 71. The spaces
between the joists are covered with boards resting on the cleats
and pushed closely together, but not nailed, so that the whole or
part of the cover can be easily removed. About 14 ft. from the
front end of the tank is a 14-in. suction pipe, L, connected with
the main suction pipe, M, which crosses all of the evaporators to
remove the water vapors. The main suction pipe, M, as well as
the branch pipes, L, are made of wooden staves kept tight by
hoops. M is connected with a large suction fan having the hous-
ing and wings of sheet copper and the shaft and arms of brass.
This fan rapidly removes the vapor from each evaporating tank,
and by its use the building, even in cold winter weather, and not-
withstanding that 1 1 such evaporators are in operation, is entirely
free from steam. A wooden stack outside the building serves
for the discharge of the fan, and the exhaust at each individual
evaporator is regulated by a wooden slide, N, inserted below
the suction pipe L.
Fig. 71 gives the construction of the first or experimental
evaporator. During the experiments it was found that the 6-in.
lead pipes passing through both ends of the tank, and being
burnt with lead tight to both the ends, did not keep their straight
position, but on account of the expansion became wavy. Profit-
ing, by this experience the back end of the other evaporators
were made sloping, and the pipes, instead of passing through,
passed over the edge of the back end. This allowed the pipes to
expand freely, and they retained their straight position.
Close to the end of the evaporating tank and resting on the
cover is the lead-lined feed-box, V, from the bottom of which is
a short pipe or nipple extending into the tank. The solution
supply pipe T, which crosses all 1 1 evaporators, is connected with
the large supply tanks, and serves to convey the solution to each
evaporator by down-takes, U. The outlet of the evaporating
tank is in the side near the furnace end, about four inches above
the hot-air pipes, B (Fig. 71).
The operation is conducted in the following manner: The pan
is first filled to the level of the outlet with the copper sulphate
solution to be concentrated, the fire is then started and the stoker
and suction fans set in motion. The big copper fan is not started
until the solution becomes hot enough to generate steam. Some
solution is added through V to keep the surface of the solution at
274 HYDROMETALLURGY OF SILVER
the same level. When it is found that the solution near the out-
let has attained the desired concentration, a continuous stream of
refined solution is allowed to flow into the tank from the feed-
box, which starts a continuous outflow of concentrated solution
through the outlet. The amount of influx is regulated by fre-
quent hydrometer tests of the solution at the outlet. The supply
of fuel by the automatic stoker being regular, the heat of the
evaporator is very uniform, and once having adjusted the proper
influx of the weak solution the outflowing stream will be found
of quite constant concentration.
The glowing hot gases entering the tubes give off the main
part of their heat to the solution within a comparatively short
distance from the point of entrance, and cause this portion of
the solution to boil. In the passage of the gases through the
tubes they gradually come into cooler regions, and are offered an
excellent opportunity to give off more of their heat to the sur-
rounding solution, so that, when they finally leave the tubes,
their temperature is much below the boiling-point of the solution;
in fact, so low that the pipes at that end can be comfortably
touched with the hand. In a tank 100 ft. or 125 ft. long the
gases would leave at a temperature about that of the surround-
ing air, thus completely utilizing the heat of the gases.
The greater economy and efficiency of this type of evapora-
tor, as compared with one having a steam coil or bottom fire, is
apparent. The production of steam involves a considerable
waste of heat, and in using it to evaporate liquids, its circulation
through coils produces a large amount of condensed water at a
temperature very nearly 100 deg. C., the heat of which is gen-
erally lost. To evaporate by direct fire under the bottom of a
pan is very inefficient and wasteful; and in the present case, in
which no other metal but lead can be used for the pan, it requires
great care and watchfulness to avoid melting the metal.
For concentrating chemical solutions which do not affect iron the
evaporator described can be constructed entirely of iron or steel,
and at much less cost. The pan itself, however, should always
be placed in a wooden tank to prevent loss of heat by radiation.
Very fine particles of ashes settle in the longitudinal tubes,
but these ashes are very light, and by turning the valve of the
compressed-air pipe E, one of which is attached to each tube,
they are easily removed and blown into the chamber 0.
EXTRACTION WITH SULPHURIC ACID 275
The continually outflowing concentrated solution passes into
a lead pipe common to all the 1 1 evaporators and is conveyed into
a collecting tank, from which the liquor is elevated by means of a
horizontal pressure tank into a system of troughs which pass
over all the crystallizing tanks, of which there are 112, each of
720 cu. ft. capacity. The troughs are covered and so arranged
that any individual tank can be filled. Wooden tanks, lead-
lined, of the same size, did not answer. By the frequent changes
of the temperature to which they were exposed, the lead lining
continued to expand without contracting again, which caused in
course of time so many leakages that it became intolerable.
The present tanks are made of 20-in. thick concrete walls, which,
however, do not come up to expectation either, though they are
far superior to the wooden lead-lined tanks. By the sudden
change in the temperature when the tank is filled with such a
large volume of hot liquor, fine cracks in the walls are caused,
through which, if not attended to, leakage will take place, but
leakage can be prevented by plastering a little cement on the
outside of the tank. An experimental tank built of bricks, how-
ever, answered the requirements of such large crystallizing tanks.
The brick walls have in the center a 2-in. space filled with a
mixture of asphaltum and sand, which combines with the bottom
layer of the tank, thus practically forming a tank by itself, embed-
ded in the brick work. When heated by the sudden filling of the
tank with hot solution, the asphaltum softens, and when gradu-
ally cooled contracts without cracking.
On the top of each tank are movable wooden frames support-
ing numerous strips of lead 5 ft. long, on which the crystals
form, as well as on the sides and bottom. The solution remains
seven days in the tank. In discharging, the mother liquor is
drawn off through a brass tube near the bottom, then the frames
with the strips are lifted up by means of block and tackle at-
tached to an overhead crawl. The crystals are knocked off from
the strips with a wooden paddle and fall into the tank. The
frames are then moved to one side by means of the crawl. The
crystals from the sides are broken down also. The tanks are
arranged in long rows intersected by several cross passages.
Between each two rows is a track for transporting the blue vitriol
crystals, on either side of which is a cement channel to receive
and to convey the mother liquor to pressure tanks for further
276
HYDROMETALLURGY OF SILVER
handling. Rails are laid in recesses near the rim of the tanks,
so that the rails of the two opposite rows of tanks form a trac"k
for a hopper mounted on wheels. The crystals are shoveled with
copper shovels into this hopper, which fills the push-car under-
neath through a spout with slide in the center. As the tanks are
6 ft. deep the crystals have to be thrown at least 7 ft., and in
doing so some of the crystals unavoidably fall back into the
tanks, frequently striking the shoveler. It was found that this
was very injurious to the men, especially in the summer. Their
bodies became covered with deep sores, which, while not danger-
ous, were very painful.
FIG. 73. — DEVICE FOR DISCHARGING BLUE VITRIOL.
To protect the men, I constructed and set in operation
the device illustrated in Fig. 73. On top of the movable hopper
is mounted the frame K, K, with a turntable and circular track
underneath, which rests on a number of stationary wheels and is
kept in place by the pin P. The object of this turntable is to
make the apparatus available for the opposite row of tanks by
rotation through 180 deg. The frame K is provided with a belt
elevator E, with copper cups.- On the shaft F are the pulleys
L and M, which drive the elevator pulley N. The elevator can
be brought to a horizontal position by means of the shaft F. The
lower end of the elevator is provided with the boot B, which can
be brought down to within a short distance from the bottom of
EXTRACTION WITH SULPHURIC ACID 277
the tank and into which the crystals are shoveled. The power is
imparted by an electric motor on the platform of the frame,
which receives the electric current by an overhead wire and
trolley. From the time this device came into operation the
men were protected, as they had to shovel the crystals only a
short distance above the floor.
The crystals are washed in a stream of mother liquor in a
trough, and conveyed by this stream into a hexagonal revolving
screen, having shaft and arms of brass. The screen itself is of
maple wood, perforated. There are two such screens, to make
two sizes of crystals. The mother liquor with the smallest crys-
tals, dirt and sediment, after leaving the second screen, is con-
veyed to an agitating tank and heated by a steam jet to dissolve
the very fine crystals. The resultant solution is sent through
a filter-press, the clean liquor flowing to the storage tanks. The
crystals are dried in ten brass centrifugal machines. The yearly
production is about 18,000 tons of blue vitriol.
The crystals having been obtained from such a pure neutral
solution are of a very deep blue permanent color, which is not
affected by light, except in the direct rays of the sun. They do
not change into a bluish white powder, which is the case with
crystals made from an acid solution.
The trade in blue vitriol demands large crystals. In crystal-
lizing a salt solution in large tanks it will be found that, while on
the strips and sides large crystals are formed, the bottom will con-
tain mostly small crystals. As this necessitated the dissolving
and recrystallization of a large portion of the bottom crystals,
and consequently of quite a percentage of the total production,
this peculiarity was rather annoying, and, searching for the cause,
I observed that on the surface of the cooling liquor numerous
very small crystals were formed. They do not remain on the
surface, but sink as soon as they are formed. These crystals are
so small that they cannot be seen as such, but if a good light
strikes the surface, the liquor right under the surface sparkles
from light reflected on these minute crystals. It can easily be
observed that "hey sink and that new ones are continually formed,
thus producing a very shower of fine crystals from the surface
to the bottom. This phenomenon is caused by the evaporation of
the water on the very surface where the liquor is in contact with
the air. By losing part of its water a very thin sheet of the
278 HYDROMETALLURGY OF SILVER
solution on the very surface will become so concentrated that it
has to form and drop these very fine crystals, aided by the cool-
ing effect of evaporation and contact with the air. Following
up these observations I caused a stream of water to enter the
tank under light pressure and level with the surface through a
flat muzzle, so that the water did not mix with the solution, but
covered as such the whole surface about an inch thick. This
stopped the sparkling of these minute crystals entirely, and when
crystallization was finished it was found that the bottom crystals
were just as good as the crystals on the sides, so that they could
be mixed and treated together with the other crystals, and the
tedious and expensive operation of dissolving and recrystallizing
of an already finished product was entirely avoided. After this
successful trial the whole crystallizing plant was equipped with
a proper arrangement for this purpose, so that the operator had
only to fill each tank to the given mark and then to turn on
the water for a short time. The mother liquor is somewhat di-
luted by this method, but the advantage gained greatly outweighs
this disadvantage.
(2) EXTRACTION OF SILVER FROM BLACK COPPER
This process is based on the reaction which takes place if
copper is moistened with warm diluted sulphuric acid in presence
of air; the latter will oxidize the copper, and the acid will combine
with the oxide, forming cupric sulphate.
This method originated and is still in operation at Oker,
Germany, for which reason it is also called the Oker process.
The material subjected to this process is argentiferous and
auriferous black copper. This is smelted in a reverberatory
furnace, being partly refined and then granulated.
The granulated copper is charged into wooden lead-lined tubs
5 ft. high, with a bottom diameter of 3J ft. and a rim diameter of
2J ft. About 5 in. above the bottom there is a movable wooden
filter bottom. Below this bottom an opening is cut out 4 in.
high and 8 in. wide, to which is attached a lead trough. This
opening is cut so large because it serves not only as outlet
for the solution, but acts also as an inlet for air. The boards
forming the filter bottom are perforated with inch holes. On
top of the filter bottom large pieces of copper are placed first,
then about one ton of granulated copper is charged on top of
EXTRACTION WITH SULPHURIC ACID 279
the coarse pieces. This done, a spray of hot diluted sulphuric
acid of 28 deg. B. (heated to 70 deg. C.) is made to play over
the copper granules; this has to be done over the whole surface.
The first acid charge will leave the dissolving tank colorless, with-
out containing much, if any, cupric sulphate. The spray of acid
is applied only for a short time. By the evaporation of the warm
solution and condensation of the steam a gentle draft is produced
through the opening under the filter bottom to the top of the
tub, and the oxidation of the copper and the formation of cupric
sulphate takes place. After a quarter of an hour or so some more
hot acid is sprayed over the granules. The stream of acid dis-
solves the cupric sulphate that is formed, leaving the surface of the
granules clean. The entering air oxidizes again the surface of the
copper, which is again washed out by acid. This is repeated over
and over again, and in fact constitutes the process. Instead of
pure acid, crude copper solution, or acid mother liquor, is mixed
with acid instead of water, which has a much more energetic dis-
solving action on the copper. The strength of such a solution is
regulated at 32 to 34 deg. B.
This is undoubtedly a very slow and tedious process. The
oxidizing part of the process would be much hastened if, instead
of depending on the natural draft of a tub 5 ft. high, an artificial
gentle stream of warm air could be forced through the tub, which
instead of 5 ft. in hight might be made 15 or 16 ft. high and
charged with three to four times the amount of granulated copper.
From time to time the tower could be flushed with cupric
sulphate solution to prevent the granules from being clogged by
the residue slimes.
Silver, gold, antimony, arsenic (the last coming mostly
from the crude acid) and lead remain as slimy residues which are
brought out from the dissolving vat with every charge of acid,
and settle in a horizontal trough 160 ft. long, 30 in. wide and 14
in. deep. The solution in this horizontal trough moves but
slowly, and when it reaches the outlet has cooled to a tempera-
ture only about 2 deg. above the temperature of the surrounding
air. The crystals of blue vitriol which form in the trough are
shoveled out on an inclined bench which lets the adhering solu-
tion drop back into the trough. This is done every three days.
The mother liquor flows into a pressure tank made of 3-in. wooden
staves and lined with 1-in. lead, which is well hooped. Instead of
280 HYDROMETALLURGY OF SILVER
compressed air steam is used, and the liquor is lifted to a reservoir
which is placed on a floor above the dissolving tanks. This
liquor contains much acid, and after mixing it with still more
acid it is heated and used for dissolving the granulated copper.
To the crude vitriol which was shoveled out from the trough
some silver slimes are adhering..- They are first washed with a
spray of water, then dissolved in a lead pan by boiling with water
or with a mixture of water and mother liquor from the second
crystallization. Crude vitriol is added until the solution measures
29 deg. B. Then the fire is stopped and the sediment allowed to
settle. When the temperature has cooled somewhat and the
solution become clear, it is very carefully decanted, so as not to
draw off any of the silver slimes, and conveyed to wooden lead-
lined crystallizing tanks of 147 cu. ft. capacity. Ten days is
allowed for crystallization.
The silver slimes consist of:
Silver 3.068 per cent.
Metallic copper 7.400 per cent.
Lead 23.100 per cent.
Lime 8.300 per cent.
Sulphuric acid 16.200 per cent.
Antimony and arsenic 27.000 per cent.
The large percentage of antimony and arsenic comes from
the impure chamber acid which is used for dissolving the copper.
These slimes are well washed and mixed with litharge while still
wet, then dried and melted; the resulting lead contains 1£ per
cent, silver.
XXI
THE ZIERVOGEL PROCESS
IN this process hot water is used as solvent for the silver,
and the latter has to be first converted into sulphate, while the
other metal sulphides, copper and iron, have to be converted into
oxides. The only suitable material is copper matte containing a
certain percentage of iron. If the matte contains too much or
too little iron the extraction of silver becomes inferior. Really
good results were only obtained from the copper matte of Mans-
feld, which contains 80 percent, copper sulphide, 11 per cent, iron
sulphide and 0.4 per cent, silver.
The great sensitiveness of this process, and the complicated
roasting it requires, limit its application. The roasting was
described in Chapter X, on sulphating roasting, and we have,
therefore, only to consider the extraction of the silver from the
roasted matte. The roasted material is charged in lots of 500
Ib. in small tubs 25 in. in diameter and 24 in. high. ^These tubs
are provided with a filter bottom, and a number of them are placed
in one row. The leaching is done with hot water of about 85 deg.
C. The water passing through the roasted matte dissolves the
silver sulphate, and leaves the tub through an outlet pipe below
the filter bottom. It flows into a square settling-tank, which
extends the whole length of the row and which receives the silver
solution from all the leaching-tubs. This tank is 30 ft. long, 2 ft.
wide and 1J ft. high, and has longitudinally a partition wiiich is
lower than the rim of the tank, and over which the silver solu-
tion flows into the second half of the tank when the first one is
filled. This tank serves to catch any material that may be car-
ried out by the stream from the leaching-tubs. Below this set-
tling-tank is placed a row of tubs 21 in. in diameter and 20 in.
deep, corresponding in number to that of the leaching-tubs. Each
of them is provided with a filter bottom. They serve for precipi-
281
282 HYDROMETALLURGY OF SILVER
tating the silver from the solution, and are filled with copper
bars 1 in. thick, 5 in. wide and 14 in. long. In these tubs nearly
all the silver is precipitated as cement silver. Flowing out from
under the filter bottom the solution enters a lead-lined box 15 in.
wide and 6 in. deep, which extends in front of all the precipitation
filters. The bottom of this box, which has several outlets, is
covered with small pieces of copper. Under each outlet is placed
a tub with filter bottom on which bar copper and granulated
copper are placed, and through which the solution from the out-
lets passes. In these tubs but very little silver is precipitated, and
the solution when leaving them is free from silver. If the water
used for dissolving the silver from the roasted matte is slightly
acidified it hastens the dissolving and causes better extraction;
it prevents the separation of basic salts.
The cement silver contains some metallic copper and some
gypsum. For purification, it is placed in tubs and rubbed with
wooden pestles to free the copper from the adhering silver, then
washed in hand pans to separate the coarser copper, next placed
in tubs and digested with diluted sulphuric acid during six to
seven days to remove the copper and as much as possible of the
gypsum, and finally it is washed with hot water. The washed
cement silver contains in 1000 parts 860 to 870 parts of fine silver.
It is dried and smelted. The desilverized solution is made to pass
over scrap iron to precipitate any copper that may have dis-
solved from the matte by the acidified water. Usually, however,
it is mixed with water and used over again a number of times
to dissolve silver instead of using pure water for the latter
purpose.
XXII
TREATMENT OF SILVER ORES RICH IN GOLD
To me was given the task of working the rich silver- and gold-
bearing concentrates from the old Tarshish mine, Alpine county,
California. These concentrated sulphides contained 258 oz.
silver and over 10 oz. in gold per ton, which represents the average
assay of five months. All the gold was contained in the sulphurets;
no metallic gold could be detected. The plan of operation
was self-suggesting, viz.: to subject the material to a chlori-
dizing roasting, then to impregnate it by Plattner's method with
chlorine, and then leach, first with water for the extraction of the
gold, and afterward with sodium hyposulphite for the extraction
of the silver. When this scheme was executed it was found, how-
ever, that while a high percentage of silver could be extracted, only
about 50 per cent, of the gold was thus extractable. The reason
is not easily explained. After operating for a while in this way
and paying particular attention to the roasting without obtaining
any better result, the operations of the process were reversed.
The ore, after being roasted with salt, was leached first with water
and with sodium hyposulphite for the extraction of the silver,
and then treated with chlorine gas by Plattner's method. This
reversing of the operations had a most beneficial influence on the
result, effecting an extraction of 95 per cent, of the gold.
The operations were as follows:
(1) Roasting. — The roasting was done with 10 per cent, of
salt, the salt being added after the oxidation had progressed for
some time. Soon after the salt was added free gold could be
detected by concentrating a sample of the ore in a horn spoon.
Roasting was continued until the concentration test did not show
any undecomposed sulphurets and but very little magnetic iron,
when the concentrated part of the sample was tested with a mag-
net, while a large amount of very bright yellow gold was visible.
283
284 HYDROMETALLURGY OF SILVER
The roasted charge, after cooling, was sifted through a screen
of 10 meshes to the running inch. The fine was charged into the
filter tanks while the coarse was accumulated in a larger lot, then
crushed dry in a battery and slightly roasted.
(2) Base-Metal Leaching. — In order to prevent, or rather to
greatly diminish, the dissolving of silver chloride by the base-
metal solution, I devised and practised the following mode of
operation: Instead of applying the stream of water on top of the
ore, as is usually done, the water was made, under very slight
pressure, to enter the vat under the filter bottom and to ascend
gradually through the ore. In this way the concentrated part of
the solution which dissolves the silver chloride accumulated on
top of the ore. If, then, this concentrated solution was diluted by
a stream of water applied on top and the solution was permitted to
flow out from under the filter bottom, the silver chloride was
precipitated on and through the ore, and was dissolved again by
the subsequent leaching with sodium hyposulphite. Sufficient
room in the vat was left above the ore for this operation.
(3) Leaching the Silver. — This was done in the usual way by
applying a diluted solution of sodium hyposulphite. The result-
ing silver bullion was 957 fine.
(4) Second Leaching with Water. — After the silver was ex-
tracted the solution of sodium hyposulphite was pressed out by
water, and washing was continued until the outflowing liquid
was perfectly free from sodium hyposulphite. Such a careful
washing is necessary, because sodium hyposulphite added to a
solution of gold chloride* prevents the precipitation of the gold
by ferrous sulphate.
The desilverized and washed ore was removed from the vat
to a drying kiln, where it was left for a time till the surplus
water had evaporated. After this it was charged back into the
vat, still moist. This second handling and drying cannot be
avoided, because the ore after leaching is too wet and tightly
packed to permit a free passage of the chlorine gas.
(5) Extraction of the Gold. — After the extraction of the base-
metal chlorides and the silver the gold is left in a metallic state,
and bright and clean, permitting a very close extraction. On the
inside periphery of the vat a groove was cut into the staves from
the rim down, forming a shoulder or recess into which a tight
wooden cover fitted. The shoulder was 2J in. below the rim, so
TREATMENT OF SILVER ORES RICH IN GOLD 285
that when the 1-in. cover was put on the staves projected 1 J in.
above the cover. Around the periphery the cover was tightly
luted with clay, and then water was poured on it to about the depth
of one inch. This sheet of water kept the cover perfectly tight.
The water, however, was not poured on the cover until the gas
appeared on the surface of the ore. The cover was provided with
two IJ-in. pieces of pipe projecting about 6 in. above the cover,
and a square opening 6 x 6 in. When the chlorine gas appeared
above the ore this opening was closed with a cover luted tight
with clay and the water poured on top of the cover. When the
gas commenced to escape through the pipes in the cover they
both were closed with balls of clay. The ore was left in contact
with the chlorine gas for twelve hours, and as soon as it was
ready for the extraction these clay balls were removed and one
of the pipes was connected with the water-pipe by a hose, while the
other was connected, by means of a hose, either with another vat
already prepared for chlorination, or with the ash-pit of the
roasting furnace. This was done to utilize the surplus of
chlorine gas, and to protect the workmen from its very injurious
effect. Care was taken to place a sack, kept in place by bricks,
on top of the ore right under the water inlet, in order to prevent
the stream from working into the ore.
Chlorine was generated in a leaden gas generator heated by
steam.
The gold solution was collected in precipitation tanks and
precipitated with a solution of ferrous sulphate. Separate tanks
were used for the precipitation of the silver. There was also a
separate line of troughs for each metal, to guard against the enter-
ing of any sodium sulphite solution into the gold solution, because,
as stated above, sodium hyposulphite prevents the precipitation
of the gold by ferrous sulphate.
In working ordinary gold-bearing sulphurets by the Plattner
method, the gold solution turns jet black when the iron solution
is added, which is caused by the precipitation of the gold in metal-
lic state, but in such an extremely finely divided condition that
it assumes this color. The concentrated sulphurets from the
Tarshish mine were very variable in their silver and gold contents,
and sometimes lots were treated containing as much as $700 to
$800 per ton in gold. When such rich gold ore is chloridized,
the solution carrying out the gold is of a very lustrous yellow
286 HYDROMETALLURGY OF SILVER
color, and if ferrous sulphate solution is added red-brown clouds
are formed, which rapidly sink to the bottom. There the gold
accumulates in spongy lumps of great specific gravity, and some
of them show scales of bright gold, which under the microscope
might prove to be crystallized gold. There is but very little
more time used in leaching rich gold ore than poor, on account of
the great solubility of gold chloride in water.
The gold was well washed, dried and melted with borax, while
the silver precipitate was melted with iron and borax in graphite
crucibles.
Results. — To ascertain the working results of this method
the concentrated raw sulphurets delivered to the reduction works
each day were carefully weighed and assayed during a period of
five months. The average value of these concentrates, as men-
tioned above, was 258 oz. silver and a little over 10 oz. gold per
ton. The total value of the bullion shipped during this period,
compared with the value of the raw sulphurets worked during
this time, showed an actual extraction of silver 96 per cent, and
of gold 95 per cent.
The gold obtained was of high fineness, varying from
970/1000 to 987/1000.
XXIII
CYANIDATION OF AURIFEROUS SILVER ORES
THE cyanide process is based on the fact that gold and silver
in presence of oxygen dissolve in an alkaline solution of potas-
sium or sodium cyanide. The cyanide solution, however, dis-
solves also other metals and their sulphides, like iron, copper,
zinc, lead, antimony, arsenic, etc., but it dissolves them much
slower, which condition makes the process possible, because, in
dissolving, these substances decompose the potassium cyanide,
and as they as a rule offer a very much larger surface than the
gold and silver, they would at equal dissolving ratio consume
so much potassium cyanide as to render the process financially
impracticable.
Some of the above-named metals and their compounds act
more energetically on the potassium cyanide than do others,
but as they all act deterioratingly it is apparent that ores
heavily charged with sulphureted minerals are not suitable for
this process unless they are first subjected to a chloridizing roast-
ing. This, of course, reduces greatly the applicability of the
process for silver ores, because much the larger part of them
are complex sulphureted ores. If the ore has to be roasted with
salt, the process enters into competition with the lixiviation
process with sodium hyposulphite, and then it is doubtful whether
it will prove to be superior, except when the ore contains a suf-
ficient amount of gold.
TREATMENT OF RAW ORE
To be suitable for this treatment the ore has to be but slightly
mineralized and the silver ore to occur in its purer varieties, as
sulphide, chloride or chlorobromide. But even from such clean
ores the extraction percentage of the silver varies greatly. Some
of them yield 90 per cent, and more of their silver to the solvent,
287
288 HYDROMETALLURGY OF SILVER
while others of similar character will yield only 50 per cent.
Much more reliable and uniform results are obtained with regard
to the extraction of the gold, for which reason the process is more
suitable for treating clean auriferous silver ores than silver ores
which do not contain any gold. Large quantities of ores can be
cheaply treated by this process in simple appliances requiring
but little repair, which give the means of working low-grade ores
which by no other known process could be worked profitably.
A number of mining properties are worked now with profit which
formerly were not productive because the ore was too poor in silver
and gold to be treated by other more costly processes.
Raw ores when finely pulverized permit only a very slow per-
colation of solution, owing to exceedingly fine slimes which are
formed in pulverizing, and which, in some cases, pack so as to be
practically impenetrable. To overcome this difficulty, the ore is
crushed wet, and the slimes are separated from the sand either by
cones and other sand-separating appliances or by conveying the
pulp from the stamp battery direct, or, where concentration is
practised, from the discharge of the concentrating tables to large
masonry tanks to retain the sand, and the overflow of these tanks
to another series of tanks to collect the slimes. The latter is a
rather crude method, permitting only a very imperfect separation,
and necessitates considerable handling of the material, while by
the former method the filling of the tanks is done automatically.
However, there may be circumstances which make the adoption
of the second method necessary or even more practicable.
In the following I give an abstract of a very careful, intelli-
gent and exhaustive record of the cyaniding of auriferous silver
ores of Palmarejo, Chihuahua, Mexico, written by T. H. Oxnam,
mining engineer, Palmarejo & Mexican Gold Fields, Ltd./
Chinipas, and read at the Washington meeting, May, 1905, of
the American Institute of Mining Engineers:
CYANIDING AURIFEROUS SILVER ORES AT PALMAREJO,
MEXICO
The predominating value of the ore now treated by the Pal-
marejo and Mexican Gold Fields, Ltd., is silver. The method
consists of wet-crushing and concentrating, followed by cyanida-
tion of the unroasted sand and slime.
The Palmarejo mines are located in the southwestern part of
CYANIDATION OF AURIFEROUS SILVER ORES 289
Chihuahua, on the foothills of the Sierra Madre, and at an eleva-
tion of 3200 ft. The mills, 12 miles distant, are situated on the
Chinipas river, near the town of Chinipas, which is 150 miles
northeast of Agiabampo, on the Gulf of California. Supplies are
shipped via this port.
M ill and Cyanide Plant. — The 50-stamp mill and cyanide
plant are situated on the Chinipas river, 1.5 miles east of Chinipas,
at a place known locally as " El Zapote." Water-power, furnished
by the river, is used to run the mill, slime plant and machine-shop.
A masonry conduit 11 miles long conducts the water to a pen-
stock a short distance above the mill, thence through a steel
pipe, 1100 ft. long, tapering from 48 in. in diameter at the pen-
stock to 22 in. at the wheel-pits; here there are four 6-ft. Pelton
wheels under a 97.5-ft. head.
The ore consists essentially of a silicious matrix in which is
disseminated a small percentage of pyrite. Black manganese
oxide, and calcite are present in varying proportions, and small
quantities of antimony and arsenic, together with traces of bis-
muth, also occur. Occasional traces of copper and zinc are found.
The major portion of the silver occurs in the form of argentite,
though a certain amount of stephanite is present, and occasionally
small patches of chlorobromide and native silver.
System of Milling. — The ore, averaging 6 per cent, moisture,
is brought to the mill in trains of from 9 to 14 cars, and is dumped
into the main upper storage-bin, which has a capacity of 1100
tons. From this the ore is drawn over 3.5 x 10-ft. iron grizzlies
having 1.5-in. openings to the 7 x 10-in. Blake rock-crushers,
which run at 250 r. p. m. and crush to 2-in. size. Of the dump-
ore, which is coarse and extremely hard, approximately 90 per
cent, goes to the crushers; of the mine-ore, which is finer and
softer, approximately 50 per cent, goes to the crushers, the other
10 and 50 per cent., respectively, falling through the grizzlies.
A secondary storage-bin (of 1100-ton) receives the ore from
both grizzlies and crushers. The ore is then trammed to three
small intermediate bins, each of 50 tons capacity; from here
it is conveyed, by means of half-ton cars, to the hoppers of the
Challenge ore-feeders. This double handling of the ore is incon-
venient, but is rendered necessary because of the construction
of the mill, which was originally erected for different require-
ments.
290 HYDROMETALLURGY OF SILVER
The stamps, when equipped with new shoes, weigh 850 Ib.
They drop 6 to 7 in., 100 times per min., the order of drop being
1 — 3 — 5 — 2 — 4; 20-mesh brass-wire screen, No. 26 wire, is used;
the hight of discharge is kept at 2 in. The stamp-duty is from
2.75 to 3.25 tons per twenty-four hours. The average stamp-
duty would doubtless be somewhat increased by the installation
of narrow mortars of the Homestake pattern. For some time
past, forged-steel shoes have been used in preference to the cast-
iron shoes of our own make. The steel shoes cost 15c. per ton of
ore crushed, as compared with 18c. for the cast-iron shoes. We
cast all our own dies, for which purpose the worn-out shoes and
iron and steel scrap are employed. The average life of a forged-
steel shoe is 95 days, while that of a cast-iron die is 33 days.
From the batteries the pulp passes directly over 10 Wilfley
concentrators, running with a |-in. stroke at 215 strokes per min.
During the year ending July 1, 1904, the concentrator saved
0.76 per cent, by weight of the ore as a product which contained
18.28 per cent, of the gold and 17.98 per cent, of the silver of the
ore.
A wooden launder conveys the pulp from the tables to the
tailing elevator-wheel. The latter is 14 ft. in diameter and is of
the outside-bucket type, having 22 steel buckets (each 18 in.
long, 8.5 in. wide and 8.5 in. deep, with a capacity of 1025 cu. in.).
The wheel is driven by a f-in. plow-steel wire cable at a speed of
18 r. p. m. The discharge efficiency, as in all wheels of this type,
is not high, the tailing leaving the wheel in a launder 5.5 ft. above
the level of the mill launder supplying the pulp.
A large masonry sand -retaining tank (divided into four
compartments, each compartment measuring 25 x 80 x 4 ft.
in depth) receives the product from the wheel. Distribution is
effected by a central launder in each compartment, provided
with a number of 4-in. side-discharge pipes. Each compartment
is provided with a removable end-discharge gate, 4 ft. wide, com-
posed of pieces of 2-in. plank, planed smooth on the edges and
sliding in guides secured to the side-posts. As the compartment
fills up with sand, the discharge of these gates is raised. The
discharge overflow empties into the main slime launder. Each
compartment also communicates with its immediate neighbors
by small side-discharge doors. The purpose of this arrangement
is that the mill product may be emptying into one compartment,
CYANIDATION OF AURIFEROUS SILVER ORES 291
from which a portion of the finer material escapes through one
of the side-gates to an adjoining compartment, while the finest
material is passing off in the discharge over the lowered end gate
of this second compartment.
It is found, however, that a considerable quantity of the finest
material will always tend to collect at the lower end of the first
compartment, receiving the discharge of the elevator-wheel, to
lessen which an overflow from the end gate of this compartment
is also necessary. The first five or six tons of material removed
from the compartments is always slimy, and is trammed a short
distance to an open drying-patio, where it is spread out, sun-
dried and broken up; after this it is mixed in with the coarser
sand and treated in the leaching- vats. A third compartment of
the sand-retaining tank is kept full of sand, which is being drained
while the dry sand is trammed from the fourth compartment.
Each of these compartments holds the sand of forty-eight to sixty
hours' crushing in the mill. The retained sand is usually sub-
jected to two days' draining before charging it into the leaching-
vats. The fine material escaping in the overflow from the masonry
retaining-tank is carried by a wooden launder to three so-called
"slime pits/' having an aggregate capacity of 15,000 tons. Every
precaution is exercised that no slime escapes at the overflow
gates of these pits; but at no time is such overflow perfectly free
from suspended matter. During the 18 months (ending Decem-
ber 31, 1904) of the total net tonnage crushed in the mill, 19.16
per cent, went to the slime pits. Sizing tests (using the ordinary
brass-wire assay screens) have shown that about 6 per cent, of
this material is retained on the 100- mesh, while 85 per cent, passes
a 200-mesh screen.
Although this material is chiefly slime (which on long drying
cracks up into layers almost impervious to leaching), yet it is
found that considerable fine but leachable sand is deposited at
the head of the slime pits and near the discharge from the slime
launder. About two months after ceasing to discharge into any
one slime pit, this fine sand at the head of the pits dries sufficiently
(during ordinary weather) to permit of being walked on; it is
then conveyed, by contract labor, to the open drying-floor, or
patio, together with a certain percentage of more slimy material
which unavoidably becomes mixed with it. Here it is spread
out, sun-dried and thoroughly broken up, after which it is
292 HYDROMETALLURGY OF SILVER
mixed in with the ordinary sand and treated by leaching. Dur-
ing the past year (1904) 2400 tons of very fine material from the
slime pits has been so treated. By far the greater portion of the
material collected in the slime pits, however, is so extremely fine
and has such a clayey nature that it is almost impervious to leach-
ing. This portion, slime, is allowed to dry as much as practicable,
and is then treated by agitation in a separate plant, as will be
described further on in this paper. Figs. 74 and 75 show the
arrangement of the cyanide leaching plant.
Cyanidation of Sand. — The sand caught in the large masonry
sand-retaining tank (after being allowed to drain as long as pos-
sible, usually from thirty-six to forty-eight hours) is trammed
in half-ton cars to the cyanide leaching- vats; these are 12 in num-
ber, 30 ft. in diameter and 4.5 ft. deep. The filter bottom (which
reduces the available depth to 4 ft. 2 in.) consists of a wooden,
lattice framework, covered by a layer of cocoa matting, over
which is stretched a filter cloth of 8-oz. duck. Two heavier
grades of duck have been tried, but they reduced the rate of
leaching and gave less satisfactory service than the 8-oz. cloth;
10 of the leaching- vats are constructed of No. 9 sheet steel; the
other two were built on the premises of 3-in. native pine. Two
additional vats, of the same dimensions and capacity, made of
3-in. redwood throughout, are in course of erection.
The sand, as charged into the leaching-vats, carries 14 to 16
per cent, of moisture; each vat holds 100 tons of dry sand. While
being trammed to the leaching-vats, slaked lime is added to each
car, and in the proportion of 4 to 5 Ib. of lime per ton. The vats
are filled and discharged by contract, for $19 a vat, equivalent to
$0.19 per ton.
Two stock solutions are employed: the weak, of 0.25 to 0.30
per cent. KCN; the strong, of 0.75 to 0.80 per cent. KCN. The
working strength of the solutions is always taken as that indicated
by titration with silver nitrate (in presence of a few drops of a
10 per cent, solution of potassium iodide, as an indicator), 10 c.c.
of the cyanide solution being taken for titration. For convenience,
we still express the strength of our working solutions in terms of
potassium cyanide, although for over a year past we have been
employing sodium cyanide exclusively. Titration with silver
nitrate shows that the sodium cyanide used is equivalent to about
125 per cent, of potassium cyanide. Our experience with sodium
CYANIDATION OF AURIFEROUS SILVER ORES 293
FIG. 74. — CYANIDE LEACHING PLANT,
PLAN.
294
HYDROMETALLURGY OF SILVER
CYANIDATION OF AURIFEROUS SILVER ORES 295
cyanide leads us to believe that it is fully as efficient as potassium
cyanide. It also appears that, since commencing the exclusive
use of sodium cyanide, our solutions become less fouled than pre-
viously. By the adoption of sodium cyanide a saving of 20 per
cent, of the freighting expense on this article has been effected.
Besides the saving in transportation expenses, the sodium cyanide
appears to possess other advantages. Other things being equal,
it would seem preferable to use a salt as pure as can be obtained.
Absolutely pure sodium cyanide is equivalent to about 132 per
cent, of potassium cyanide; a product, testing from 125 to 130
per cent, of potassium cyanide, is nearly pure. It by no means
follows, however, that the ordinary commercial cyanide, rated as
98 to 99 per cent, pure, contains but 1 to 2 per cent, of impurities.
That this commercial cyanide frequently carries a varying per-
centage of sodium cyanide is a well-known fact; and it of course
naturally follows that the greater the percentage of sodium cya-
nide contained in the ordinary 98 to 99 per cent, potassium cyanide,
the greater the percentage of impurities.
As soon as a vat is charged, from 20 to 25 tons of weak solu-
tion (carrying, as just stated, from 0.25 to 0.30 per cent, of KCN)
is introduced, from the bottom, by means of a 2-in. drop-pipe,
terminating in a T underneath the filter. This solution is in-
troduced slowly in order to avoid channeling of the charge; it
usually makes its appearance on top of the sand about six or
seven hours after being turned on. When the solution stands 2
or 3 in. above the top of the charge, it is turned off, and the
material is allowed to soak for six hours, during which time the
sand will usually have settled from 3 to 4 in. The weak-solution
discharge-valve at the bottom of the vat is now opened and
leaching is commenced. During the next two or three days
weak solution is added from the top, as rapidly as permitted by
the leaching rate of the charge, until a total of from 100 to 130
tons has been applied. From 60 to 70 tons of strong solution,
averaging between 0.75 and 0.80 per cent, of KCN, is now run
through the charge at a somewhat slower rate, the usual time
consumed by this operation being forty-eight hours. Weak solu-
tion is next run through the charge as rapidly as possible, until
twenty-four hours before the time it is to be discharged; then
wash-water, to the amount of 15 to 20 tons, is added in lots of 5
tons each. The residue is then ready for sluicing, which is accom-
296 HYDROMETALLURGY OF SILVER
plished by two men in about six hours, each using a 2-in. hose,
equipped with a 0.5-in. nozzle and operating under a head of 72
ft. After finishing the sluicing, the canvas filter is usually swept
clean with a broom; if this is not done, it is found that the filter
cloth clogs with fine slime and the rate of filtration is lowered.
Each vat is equipped with two 10 x 10-in. square bottom-
discharge doors.
The quantity of wash-water used is regulated by the balance
of the solutions on hand. Although a separate zinc-box is pro-
vided for waste solution, it is seldom used except during the rainy
season; during other parts of the year but little solution is run
to waste. Only two of the leaching-vats are under cover, and
during the rainy season it becomes necessary to run a certain
percentage of the solution to waste; each heavy rain gives the
exposed vats a very appreciable quantity of water.
It has been my experience that thorough oxygenation of the
material is a very desirable feature in the cyanidation of gold
ores; in the case of the Palmarejo ores, this is essential to obtain
the best results. Due to the fact that the major portion of the
value is in silver, the actual weight of fine metal to be acted upon
is much greater than is ordinarily the case with gold ores.
In order to permit as much air as possible to be supplied to
the sand during treatment, the solution is frequently allowed to
drain down several inches beneath the surface of the charge; air
is thus allowed to penetrate the material to this depth. It is our
custom to assay each charge every twenty-four hours, after the
first five days of treatment. Before each sampling, the solution is
allowed to drain down several inches below the surface sand,
thus allowing additional opportunity for the entrance of air into
the upper layer of the charge.
Under the most favorable conditions, however, the air drawn
into the top layer can have but little effect on the lower half.
It is doubtless due to this difference in aeration of the upper and
lower portions that the lower half of the tailing will run from 1
to 2 oz. of silver higher than the upper half. Frequently this
difference is even more marked, a variation of 3 or 4 oz. being
obtained between the upper foot and the bottom foot of the
residue.
To overcome this, and after many experiments, some time ago
the practice was adopted of transferring as many charges as pos-
CYANIDATION OF AURIFEROUS SILVER ORES 297
sible from one vat to another during the treatment. To transfer
a vat means the loss of practically twenty-four hours of its avail-
able leaching time, because it is necessary to drain the charge
for twelve hours before commencing to transfer it. Also, it is
necessary that one of the adjoining vats be empty at the proper
time to receive the transferred charge. By careful manipulation,
at present about one-third of the total number of charges treated
are transferred. When the two additional vats, now in course of
erection, shall be completed, a greater number of charges can be
transferred and the additional capacity afforded will also permit
a longer treatment to offset the time lost in transferral. The
transferring is done by contract for $16 a vat, which is equal to
16c. per ton. While being transferred, the material is of course
given a thorough exposure to the air; any existing lumps are
broken up by the shoveling; and, roughly speaking, the bottom
layer of the original charge becomes the top layer of the trans-
ferred charge. Operations are usually so timed that the trans-
fer takes place while the strong solution is in contact with the
material. During the transfer, 100 Ib. of slaked lime is evenly
distributed near the bottom of the vat receiving the transferred
charge.
The charge is sampled just before and just after transferral,
the latter sample being 1 oz. higher in silver than the former, a
result doubtless due to the fact that the tendency is to obtain a
larger percentage of the top half of the charge than of the lower
half; and, as heretofore mentioned, the lower half of the original
charge, after the transferral, becomes practically the upper half
of the transferred charge.
The first solution added after the transfer is introduced slowly
from the bottom, after which the regular routine treatment is
continued. The value of the effluent solution from a charge is
found to increase immediately after the charge has been trans-
ferred, such increase being usually from 2 to 3 oz. of silver per
ton of solution.
In general, all head- and tailing-samples of the sand are
taken with a 1.5-in. auger at 12 to 18 different points.
The record of a single charge, which is representative of what
is regularly obtained in ordinary operations, is as follows:
Extraction, 96.27 per cent, of gold and 53.21 per cent, of
silver.
298 HYDROMETALLURGY OF SILVER
Total time of treatment, including charging and discharging,
11 days.
Solution added: weak, 261; strong, 64; wash-water, 15; total,
340 tons.
All tailing-samples, with the exception of the discharged tail-
ing, were washed before assaying.
During transferral a sample taken from upper 18 in. of charge
assayed: $0.50 of gold, 9.20 oz. of silver; and one from the lower
18 in. of charge assayed $0.82 of gold and 11.92 oz. of silver.
The effluent solution from the leaching-vats is carried to the
sump-tanks by two separate lines, one for the weak, the other for
the strong solution. These tanks are of masonry and are three in
number. Two of them (having a combined capacity of 65 tons)
are connected and serve as a weak-solution sump; the other,
having a capacity of 25 tons, is used for the strong solution.
All solution draining from the leaching-vats is passed through the
zinc-boxes before being returned to the vats.
The proper tonnage of strong solution is maintained by deter-
mining the strength of the effluent solution from the leaching-
vats; when this strength reaches 0.35 per cent, of KCN, the
solution is turned into the strong-solution sump. As a working
guide for maintaining the proper alkalinity of stock solutions,
they are titrated every day with the addition of about 5 c.c. of
strong lime-water; 10 c.c. of cyanide solution is used in all titra-
tions. If the addition of lime-water causes a difference of more
than 0.5 Ib. in the indicated strength of the solution, the quantity
of lime added to the sand charged into the leaching-vats is
increased.
From the sump the solution is elevated by a 3-in. centrifugal
pump (900 r.p.m.), to three storage-tanks at the head of
the zinc-boxes, a vertical distance of 29 ft. and a horizontal dis-
tance of 150 ft. These tanks are each 10 ft. in diameter, 8 ft.
deep, and have a capacity of 19 tons. Two of these tanks are
used for the weak, and one for the strong solution. The solution
from the vats now passes through the zinc-boxes, from which it is
led to three storage-solution tanks beneath the boxes. These
storage-tanks are made of No. 9 sheet steel, each being 15 ft. in
diameter, 6 ft. deep, and with a capacity of 33 tons. Two of
them are used as strong-solution storage-tanks; the other, as a
v eak-solution storage-tank. The strong solution is brought to
CYANIDATION OF AURIFEROUS SILVER ORES 299
the required strength by adding cyanide to the last compartment
of the strong-solution zinc-box, which is reserved for this pur-
pose. No cyanide is added directly to the weak solution.
Precipitation of Silver and Gold. — There are six zinc-boxes,
five for the weak and one for the strong solution. The five weak-
solution boxes are constructed of No. 10 sheet-steel, and are 2 ft.
wide and 18 ft. long over all. Each box contains eight compart-
ments, each compartment having an available zinc capacity of
24 x 24 x 18 in., equivalent to 6 cu. ft. Six compartments
only are filled with zinc shavings; therefore, each box, when
freshly dressed, contains 36 cu. ft. of zinc shavings, making a
total of 180 cu. ft. of zinc shavings in the five weak-solution
boxes.
The strong-solution zinc-box consists of seven individual
round boxes or compartments, placed in series, each compart-
ment being 28 in. in diameter and 24 in. in depth, and having an
available zinc capacity of 5 cu. ft. Only six of the compartments
are filled with shavings, the last compartment being reserved for
the addition of the quantity of cyanide required to bring up the
strong solution to standard strength. The strong-solution zinc-
box has, therefore, a total of 30 cu. ft. of zinc shavings.
Records are kept of the quantities of weak and strong solu-
tion daily passing through the boxes, together with their assay
values before and after precipitation. These records for the year
(1904) show that 91,793 tons of weak, and 22,251 tons of strong,
solution passed through the boxes ; this is equivalent to an aver-
age of 251 tons of weak, and 61 tons of strong solution every
twenty-four hours. During this period the flow of solution
through the boxes was interrupted on various occasions for a
short time, due to the ordinary clean-ups, dressing of the boxes
and unavoidable delays. Without taking such stoppages into
account, the average rate of flow through the boxes equaled 1.4
tons of weak solution per twenty-four hours per cubic foot of shav-
ings, and 2.03 tons of strong solution per twenty-four hours per
cubic foot of shavings.
The actual rate of flow exceeds these figures, as it is assumed
that the boxes were at all times kept dressed with the maximum
amount of shavings, which was seldom the case.
The shavings are cut on an ordinary zinc lathe, from No. 9
sheet zinc, the size of the sheets being 18 x 84 in. Ordinarily,
300 HYDROMETALLURGY OF SILVER
six sheets are wound on the mandrel of the lathe for one cutting.
One boy, working twelve hours, cuts sufficient shavings to supply
both the leaching and agitation plants, which together require an
average of 120 Ib. per twenty-four hours. It is found best to
keep only a few days' shavings on hand; freshly cut shavings give
better results than those which have been cut for some time.
The customary practice of moving the zinc from the lower to the
upper compartments, when dressing the boxes, is not followed,
fresh zinc being added as required to the top of each compart-
ment.
The strength of the solution running through the weak boxes
will average between 0.25 and 0.30 per cent, of KCN; while that
of the solution going to the strong zinc-box will average between
0.35 and 0.45 per cent, of KCN.
The average assay values per ton of the solutions entering
the zinc-boxes are approximately as follows:
Weak solution, $1 of gold and 2.25 oz. of silver; strong solu-
tion, $1.24 of gold and 3.5 oz. of silver.
It is seldom that any trouble is experienced with the precipita-
tion of the contained values. As a rule, the precipitation of the
gold is practically perfect; that of the silver averages 95 per cent.
When precipitation falls off, it is usually due to the presence of
an accumulated excess of lime in the solution.
Clean-up of Zinc-boxes. — On account of structural difficulties,
it is necessary to handle the precipitates more than is desirable.
The boxes are cleaned twice a month. Before commencing on
any box, clear water is passed through it a sufficient length of
time to displace most of the cyanide solution; this requires 10 or
15 minutes. The shavings in the first compartment are thor-
oughly washed, after which they are removed and the water bailed
out into the next compartment. The precipitates are now con-
veyed by buckets to the clean-up box, where they are passed
through a 20-mesh screen. A small percentage of "short" zinc
passes through this screen, but the greater part of such product
is here separated from the finer precipitate and is returned to the
boxes. The first compartment is now filled with water; the zinc
contained in the other compartments is gradually transferred to
it and thoroughly washed, the precipitates from each compart-
ment being carried to the clean-up box as before mentioned.
To minimize the oxidizing effect resulting from exposure of the
CYANIDATION OF AURIFEROUS SILVER ORES 301
wet zinc to the atmosphere, the washed shavings are at once
placed in the highest vacant compartment of the zinc-box and
covered with solution.
The precipitate accumulating in the first compartment from
the washing of the shavings, after settling for a short time, is
also removed to the clean-up box. This latter is provided with
three smaller settling-boxes, placed in series, which take the
overflow from it. The bottom of the clean-up box is tapped by
a 4-in. drop-pipe, which discharges directly into two large drying-
pans beneath.
The product is now dried as much as is practicable, and then
mixed, carefully sampled, assayed and sold on the premises to
one of the large ore-buying companies. The moisture in the
dried precipitate has averaged 0.27 per cent, during the past year.
The clean-ups are bulky; the net dry-weight of precipitate in
each clean-up averaged between 1100 and 1200 Ib. avoirdupois
during the past year.
Considering the fact that the precipitates receive no treatment
whatever beyond being passed through a 20-mesh screen, and the
simple drying, as above mentioned, it is rather surprising that
they carry such a high percentage of fine metal. During 1904
the assay returns, on which the sale of the precipitates is based,
have averaged slightly over 20,000 oz. silver, and approximately
$8000 of gold, per short ton. By actual weight, therefore, the
percentage of fine metal contained in the dried product recovered
throughout this period was approximately 68.57 percent, of silver
and 1.33 per cent, of gold, making a total of 69.90 per cent, of both
metals.
Two clean-ups during 1904, of a combined net weight slightly
exceeding 2300 Ib., gave an average assay value of 22,200 oz.
of silver per ton, making the fine-silver content equal to 76.12
per cent, by weight of the precipitates.
A record of the labor employed in clean-ups shows that four
men (one American and three native helpers) would readily re-
move 1200 Ib. (net dry-weight) of precipitate from the boxes,
and have the product in the drying-pans in eight hours. Based
on the average assays of the precipitate for the year, this means
that in thirty-two man-hours approximately 12,200 oz. of fine
metal would he handled, being equivalent to a duty of 371 oz.
per man-hour. This rather high duty is of course due entirely
302 HYDROMETALLURGY OF SILVER
to the fact that the precipitate contains high percentages of pre-
cious metals.
Tonnage and Extraction. — During 1904, 34,900 tons of sand
were treated in the leaching-plant. This tonnage would have
been considerably greater had it not been that during this period,
aside from the stoppages due to general repairs, the mill was closed
down for intervals aggregating 57 days for the entire 50 stamps.
The extraction for 1904 (as indicated by the assay differences
between the sand charged into the leaching-vats and that being
discharged) was 95.5 per cent, of the gold and 52.5 per cent, of the
silver value. The combined total during the year checks closely
with that called for by the sand assays. The assay value of the
sand treated during this period averaged $2.85 of gold, and
slightly more than 16 oz. of silver per ton. During 1904, the re-
turns from the precipitates were practically 1 per cent, less in gold
and 0.5 per cent, more in silver than those called for by the pre-
cipitation records.
Consumption of Cyanide, Zinc and Lime. — The office records
show that for 1904 the consumption per ton of sand cyanided was
as follows: cyanide, 2.95 lb.; zinc, 0.96 lb.; and lime, 4.33 Ib.
Expressed in terms of potassium cyanide, this consumption
would equal 3.69 lb. of potassium cyanide per ton of ore treated.
General Remarks. — The total quantity of solution passing
through the zinc-boxes during 1904, divided by the quantity of
sand for the same period, shows that, for each ton treated, 3.27
ton of solution left the leaching-vats, of which 2.63 tons was weak
and 0.64 ton strong solution. It is found that a large quantity
of weaker solution gives more satisfactory results than a small
quantity of the strong, and it is always made an important point
to pass as much weak solution through a charge as possible.
Experience has demonstrated that, in a given length of treat-
ment, a rapid leaching rate and a large quantity of solution are
more efficient than a slower leaching rate and a consequently
lesser quantity of solution. The solution pipe-lines and launders
occasionally become quite choked in places with scale deposited
from the solution. This scale, taken from lines carrying precipi-
tated solution, contained from a trace to $1 of gold and from 1
to 7 or 8 oz. of silver per ton. The scale deposited from the un-
precipitated solution usually runs higher, several assays taken
having averaged about $5 of gold and 18 oz. of silver per ton.
CYANIDATION OF AURIFEROUS SILVER ORES 303
Ordinarily the solutions do not become excessively fouled.
They contain small quantities of iron and manganese in addition
to the zinc compounds present. Alkaline sulphides are very
rarely or never noticed in solution; however, sulphocyanates and
ferrocyanides appear to be constantly present in fair quantities,
about 0.41 per cent, of ferrocyanide and 0.048 per cent, of sulpho-
cyanate.
The sand charged averaged about 0.09 per cent, of latent
acidity; and, as a rule, it contained no free acid.
The concentrate produced is sold to the same company that
buys the cyanide precipitate. An attempt had been made to
treat the concentrate by cyanide, but without success. Experi-
ments on both raw and dead-roasted concentrate (reduced to
various degrees of fineness) by leaching and agitation, for vary-
ing periods of time up to 34 days, and using solutions varying
from 0.2 to 2 per cent, of KCN, proved unsatisfactory.
Table I, given herewith, shows the working costs for milling
and cyaniding during 1904. The cost of all supplies is increased
by the heavy freight transportation expenses, as well as by the
duties placed by the Mexican Government on most of the supplies
used.
TABLE I — WORKING COSTS PER TON
Milling:
Supplies $0.640
Labor 0.357
Lubricating 0.023
Assay office (labor and supplies) • 0.035
Concentrating 7~: . . . . 0.092
Power (ditch maintenance and supplies) 0.234
Salaries 0.264
Miscellaneous (lighting, etc.) 0.018
Management and general expenses 0.336
Total * $1.999
Cyaniding:
Cyanide (2.95 Ib. @ $0.63) $1.859
Zinc (0.96 Ib. @ $0.30) 0.288
Lime (4.33 Ib. @ $0.0118) 0.051
Other supplies 0.050
Labor . ... 0.329
Salaries 0.371
Assay office (labor and supplies) 0.036
Power (ditch maintenance and supplies) 0.017
Miscellaneous (lighting, etc.) 0.004
Management and general expenses 0.186
Total 2 $3.191
1 $1.999 Mexican currency during this period was equivalent to $0.95 gold.
2 $3.191 Mexican currency during this period was equivalent to $1.22 gold.
304 HYDROMETALLURGY OF SILVER
The cost of realization on cyanide precipitate has not been
included in the cyanide working costs given herewith. Trans-
portation expenses on the precipitates are also very heavy.
In addition to these comes the heavy item of government bullion
taxes. The average cost of realization on cyanide precipitate,
per ton of ore cyanided, is as follows: Government taxes, $0.84;
treatment charges (including transportation expenses), $1.06;
total, $1.90. The cost of realization on the concentrate produced
is also unusually high, on account of the heavy transportation ex-
penses and government bullion taxes. The average cost of
realization per ton of ore crushed is as follows: Government taxes,
$0.35; treatment charges (including transportation expenses),
$1.08; total, $1.43.
Treatment of Slime. — As already mentioned, the accumulated
and currently produced slime is now being treated in a separate
plant by a system of agitation and decantation, centrifugal
pumps being used as the means of agitation. The slime plant
consists essentially of the following parts and accessories:
Four agitation and four decantation vats, each provided with
conical bottoms and connected with its own separate centrifugal
pump; two solution tanks placed at the head of the zinc-boxes,
which receive the solution from the decantation vats; four sets
of zinc-boxes and three solution sumps, which receive the solu-
tion leaving the zinc-boxes; one special solution tank, placed
at a higher level than the rest of the plant, and used prin-
cipally to supply solution to the pump bearings under pressure;
two ordinary 3-in. centrifugal pumps, used only for pumping
solution from the sumps to any desired vat or to the upper
solution tank just mentioned, they being so connected up that
either pump can be used should the other get out of order.
Each pump is run by a friction-clutch pulley, which enables
it to be started or stopped in a moment, independently of the
other pumps. A small 14 x 15-in. friction-geared hoist is used
to convey the slime from the slime pits to the agitation vats.
The entire plant is run by a 5-ft. Pelton wheel, making about
115 r.p.m., and operating under a head of 81 ft., using a 4-in.
nozzle. Water-power is obtained by means of a 14-in. riveted
steel pipe, tapping the main pipe line supplying power to the
mill. This 14-in. pipe-line was brought in by mule back, riveted
in 10-ft. lengths, although some difficulty was experienced in its
CYANIDATION OF AURIFEROUS SILVER ORES
305
transportation. (Figs. 76 and 77 give plan and section of the
slime plant.)
The method of treating the slime is similar to that ordinarily
practised by agitation and decantation; it consists in giving a
two days' agitation in the agitation vats, with from two to three
times the weight of cyanide solution, followed by another two
(or sometimes three) days' treatment in the decantation vats;
during the latter treatment, the charge, after having been suffi-
ciently agitated with the addition of slaked lime, is allowed to
settle as much as practicable, and the clear supernatant liquor is
decanted and passed through the zinc-boxes. This operation
of agitation, settling and decantation of clear solution is repeated
as many times as permissible within the time limit of the treat-
ment, being ordinarily three or four decantations.
The material treated, when dried to 20 per cent, or 25 per cent,
moisture, is tough and of the consistency of soft putty. It con-
tains, however, a certain percentage of fine sand and, when viewed
in vertical section, presents a somewhat stratified appearance.
On long drying, it cracks into layers which are almost absolutely
impervious to leaching.
The results of the sizing-test (which are given in Table II)
represent an average obtained from the material treated.
TABLE II — SIZING-TEST ON SLIME
Assay value of material was $4.13 of gold and 20.30 oz. of silver per ton.
SIZE OF MATERIAL
WEIGHT
ASSAY VALUE
PERCENTAGE OF
TOTAL VALUES
CONTAINED
GOLD
SILVER
GOLD
SILVER
Retained on 80 mesh. . .
100 ...
120
150
200
Passed 200
Per Ct.
1.1
2.7
5.6
3.1
2.7
84.8.
$2.38
2.06
1.96
2.27
2.16
4.54
Oz.
14.22
13.60
13.02
14.14
13.10
21.68
Per Ct.
0.83
1.35
2.66
1.70
1.41
93.22
Per Ct.
0.77
1.81
3.59
2.16
1.74
90.52
Totals
100.0
100.92
100.59
Description of the Plant. — The four agitation vats, constructed
of 3-in. redwood throughout, are provided with conical bottoms,
slanting at 45 deg. As shown in Fig. 78, each vat has an inside
diameter of 15 ft. 7 in.; the vertical depth, from top of side staves
306
HYDROMETALLURGY OF SILVER
FIG. 76. — PLAN OF SLIME PLANT.
CYANIDATION OF AURIFEROUS SILVER ORES 307
308
HYDROMETALLURGY OF SILVER
to the iron casting at point of conical bottom, is 15 ft., the inside
depth of vertical side staves being 7 ft. 3 in. Each agitation vat
is connected with a special manganese-steel lined 4-in. centrifugal
pump, which runs at a speed of 900 r.p.m. The pump is con-
nected with the vat by the 4-in. suction pipe a, which enters the
vat through the side staves about 6 in. above their juncture with
the bottom staves and extends nearly to the center of the vat,
where it is connected by means of a movable elbow, 6, with a
short piece of 4-in. pipe, c, provided at the free end with a good-
sized screen or strainer, d, made of J-in. sheet iron, punched with
a number of 1-in. holes; this short piece of pipe, together with
FIG. 78. — AGITATION VAT.
the screen, is of such a length that when being lowered the screen
will just clear the bottom staves. The screen is provided with a
small iron ring, to which is fastened a piece of rope, by means
of which it can be raised and lowered.
Just outside the vat, the suction-pipe is provided with an
air-cock, e, which admits air to the material going through the
pump. This air-cock, however, is rarely used at the present
time. The service-cock / permits the shutting off of the material
from the pump at any time it may become necessary, as, for in-
stance, to repack the stuffing-box or to examine the interior of
CYANIDATION OF AURIFEROUS SILVER ORES 309
the pump. The 2-in pipe-line g, provided with the valve h, con-
nects with the upper solution tank.
When it becomes necessary to shut the pump down for any
length of time (either at the conclusion of the agitation of the
charge or at any time during the treatment), the 2-in. valve h
is opened and the clear solution only passes through the pump.
The friction-clutch pulley running the pump is now thrown out
of clutch, and after the pump has stopped the valve h is closed.
By this means is avoided the accumulation in the pump interior
of solid matter that would naturally be deposited (when the
pump is stopped for any length of time) from the slimy material
ordinarily passing through it.
The 4-in. discharge-pipe i of the pump is provided with a small
bib-nosed petcock, /, a few inches from the body of the pump,
by means of which samples can readily be taken of the material
passing through the pump. The discharge-pipe passes over the
top of the vat, and at a point vertically over the center of the
bottom casting is provided with an elbow and drop-pipe, k, which
reaches to within about 15 in. of the bottom casting. This pipe
is held firmly in position by means of an iron clamp and four legs
made of f-in. bolts fastened to the bottom casting and which
serve as a support. The distance of the lower end of this dis-
charge-pipe from the bottom of vat is a matter of some impor-
tance in the agitation, and a number of experiments made along
this line have indicated that the best satisfaction is obtained at a
distance of 15 in. from the bottom casting.
Different shapes of nozzles have been tried at the lower end
of the drop-pipe, but experience has shown that the plain 4-in.
pipe-end gives satisfactory results. The discharge-pipe of the
pump tends to act as a siphon when the pump is stopped at any
time during the agitation, and would therefore cause inconvenience
when repacking the stuffing-box or making any necessary repairs.
To prevent this, air is admitted to the pipe by opening the small
air-cock I, tapped into the elbow at the upper end of the drop-
pipe. This air-cock I is also frequently used to allow the air
entering into the charge to be agitated, it being found for this
purpose preferable to the air-cock on the suction-pipe. It might
be supposed that when this air-cock is open during the agitation a
steady stream of the material passing through the discharge-pipe
would be ejected through it; and with regard to the air-cocks
310 HYDROMETALLURGY OF SILVER
similarly situated on the pump connections of the decantation
vats, such is the case. As regards the pumps connected with
the agitation vats, however, the effect is found to be quite the
reverse, and a rather strong air suction usually occurs when this
air-cock is open.
The pump-bearing nearest the pump shell is tapped with a
small pipe-line, m, provided with the valve n, which connects with
the upper solution tank. By this means the bearing is supplied
with clear solution under pressure, and the wear on the shaft and
bearing is greatly reduced. At the commencement of operations
clear water was supplied to the pump bearings and was also used
for cleaning out the pumps and for priming, when necessary.
It was soon found, however, that the quantity of water added in
this way increased the volume of stock solution very appreciably,
and, of course, an equal quantity of weak cyanide solution had
ultimately to run to waste. Not only did this cause an unneces-
sary mechanical consumption of cyanide, but the quantity of
water added through the pump bearings naturally reduced the
strength of the working solution in the vat under operation, with
a deleterious effect on the percentage of extraction. The quantity
of solution that will be added to a vat during the usual period
of agitation (from forty to forty-four hours), when the shaft and
bearing are a little worn, is surprising, amounting in some cases to
15 tons, even when the greatest care is exercised. The amount of
solution added in this way is naturally the least just after the
pump has been equipped with a new shaft and new liners, and
the bearing rebabbitted. On an average, however, the quantity
of solution added to each charge through the pump bearings is
from 5 to 6 tons. The agitation pumps in use, while in most
respects proving very satisfactory, have nevertheless certain de-
fects in their design, which contribute largely to the rapid wear-
ing of the shaft and the bearing next to the pump shell, and also
to the wearing of the interior, renewable manganese-steel wear-
ing parts. The life of these parts naturally varies; but ordinarily
it is necessary to equip a pump with a new shaft and with por-
tions of the manganese-steel wearing parts, and to rebabbit
the bearing every six weeks. The pumps are equipped with a
pulley having a 6-in. face; but it is found preferable to use a
4-in. belt, since this reduces the weight on the pump-shaft with
a consequent decrease in its wear, while a 4-in. belt runs the
CYANIDATION OF AURIFEROUS SILVER ORES
311
pump equally as well as a 6-in. one. Wire lacing is used on all
the belts.
Each agitation vat was originally provided with a 6-in. dis-
charge opening at the center of the bottom casting. This open-
ing was bushed down to 4 in. and was provided with a nipple
and a straightway valve. The first few vats were discharged
from the bottom by this means, but a deal of trouble was experi-
enced, due to the fact that, though all the slime entering the
agitation vats passed through a grizzly having 1.25-in. openings,
yet the bottom valve would frequently become choked with small
rocks and other material which seemed to be mixed with the
first slime. This bottom discharge was therefore discontinued,
FIG. 79. — DECANTATION VAT
a hole was bored in the bottom staves 10 in. from the bottom
casting, and a 3.5-in. iron service-cock was secured to the vat by
means of a short nipple and iron flanges. The vats are discharged
through this valve into a wooden launder which conveys the
material to the corresponding decantation vat. This launder is
provided with rows of 6-in. wire nails, which serve to catch any
foreign matter.
The four decantation vats, made of 3-in. redwood through-
out, are of the same dimensions as the agitation vats, with the
exception that they are provided with conical bottoms, slanting
at 20 deg. Each one is connected with an ordinary 3-in. centrif-
ugal pump. Fig. 79 shows in detail the connection of the pump
312 HYDROMETALLURGY OF SILVER
with the vat. The vat is discharged through a 3.5-in. bottom-
discharge valve and pipe, into the residue launder, from which
the discharged material flows to the river. Removal of the clear
solution is effected by means of a 2-in. decantation pipe and float.
This pipe enters the side of the vat about 6 in. above the bottom
staves and is provided writh two loosely threaded elbows, which
permit of the free raising and lowering of the portion within the
vat. The float proper is made of two ordinary 5-gal. oil cans,
soldered water-tight and painted with paraffin paint. The rate
of decantation is controlled by means of the 2-in. valve just
outside of the vat.
It frequently happens that the solution drawn from the decan-
tation vats is not perfectly clear, and two filter-boxes are pro-
vided (see Figs. 76 and 77) for the partial clarification of the
solution before it enters the solution tanks at the head of the
zinc-boxes. Each compartment of these filter-boxes is provided
with a discharge valve, by means of which the sediment deposited
from the solution can be washed into a waste launder.
The solution tanks at the head of the zinc-boxes are two in
number, one being used for the weak and the other for the strong
solution. They are made of 2-in. redwood throughout, and are
each 11 ft. 8 in. in diameter and 7 ft. 7 in. deep, inside measure-
ments, having a capacity of 25 tons. Each solution tank is
provided with a 2-in. floating hose, by means of which the clearest
solution in the tanks is always supplied to the zinc-boxes. A 3-in.
valved opening in the bottom of each of these solution tanks per-
mits of the discharge of the accumulated slime into a waste
launder.
Fig. 80 shows the timber foundations supporting the decanta-
tion vats, the conical bottoms resting on three beveled rings.
The supports for the agitation vats are built in the same manner,
the supporting rings, however, being placed to line at 45 deg.
instead of at 20 deg. Fig. 81 shows the decantation vats in course
of erection.
. There are four sets of zinc-boxes, each set being composed of
six round individual boxes or compartments, each compartment
being 28 in. in diameter and 2 ft. deep, and having an available
zinc capacity of approximately 5 cu. ft. One of the boxes is
used solely for strong solution, and two for weak solution; the
fourth being so connected up that either weak or strong solution
FIG. 80. — Timber Foundations supporting Decantation Vats of Slime Plant.
FIG. 81 — Decantation Vats in Course of Construction.
CYANIDATION OF AURIFEROUS SILVER ORES 313
may be run through it. The solution leaving the zinc-boxes
passes to three sump-tanks, made of 2-in. redwood throughout,
each 11 ft. 8 in. in diameter and 9 ft. 7 in. deep, inside measure-
ments, and of a capacity of 32 tons of solution. Two of these
tanks are connected together and serve as a weak-solution sump,
the other being used for the strong solution.
Fig. 82 gives a good view of the plant shortly before its com-
pletion and shows its general arrangement. Fig. 83 gives a
nearer view of three of the agitation vats and shows the tops of
two of the decantation vats. The 4-in. centrifugal pump (lead-
ing to No. 1 agitation vat) is seen partially connected up. Pro-
truding from the top of the decantation vat, a little below the
center of the picture, is seen the end of one of the 2-in. decanta-
tion pipes.
Method of Treatment. — The accumulated slime, after having
been dried in the slime pits as much as practicable, is conveyed
to the agitation vats in ordinary half-ton ore cars by means of a
small, friction-geared hoist. Each agitation vat is provided
with an iron grizzly (measuring 3 ft. 3 in. x 9 ft., and having
1.25-in. openings), which is suspended over to one side of the
center. The content of the car is dumped on to this grizzly and
the portion that does not pass of its own weight is trampled, or
otherwise forced through, by boys. For some time the material
being treated averaged from 20 to 25 per cent, of moisture and in
this condition was lumpy and cohesive. During this period the
agitation was unsatisfactory and the percentage of extraction
was low. Difficulty was experienced in discharging the vats;
the unagitated portion of the charge would remain in the pointed
bottom of the vat as a tough, putty-like mass, after all the liquid
portion had been discharged, and could only be washed out by
means of a stream of solution or water under pressure. Experi-
ence demonstrates that the best condition of the material is such
that, when dumped on the grizzly, it will run through of its own
weight. In this state the slime carries from 30 to 35 per cent,
moisture. It is desirable that the percentage of moisture con-
tained in the slime when charged shall be as low as possible, com-
patible with satisfactory agitation; the greater the percentage of
moisture contained in the slime, the greater will be the mechanical
consumption of cyanide. The complete drying of the slime by
some cheap process, followed by powdering before charging into
314
HYDROMETALLURGY OF SILVER
the agitation vats, should be productive of improved results.
A charge equivalent to about 15 tons of dry slime gives more
satisfactory results than does a heavier one.
TABLE III. — SETTLING RATE OF SLIME PER HOUR, WITH
ADDITION OF LIME
SETTLEMENT (IN INCHES) OF SLIME
Test
No. 1
Test
No. 2
Test
No. 3
Test
No. 4
Test
No. 5
(e)
Test
No. 6
(<0
Test
No. 7
(c)
Proportion of solution to
slime
2.5:1
2.5:1
2.5:1
2.5:1
2.5:1
2.5:1
3.3:1
Lime added per ton of
slime (<z) . ...
21b.
31b.
31b.
31b.
31b.
None
(6)
41b.
1 hour
11.0
10.5
10.0
16.0
14.0
15.0
22.0
2 hours
21.0
190
165
250
21 0
24 5
365
3
27.5
26.0
23.5
33.0
300
335
51 5
4
33.0
32.0
30.0
40.0
390
400
54 0
5
36.0
35.5
36.0
42.0
43.0
42.0
570
6
7
38.0
39.5
38.5
40.0
40.0
41.5
43.0
44.0
47.0
48.5
43.0
44.0
58.0
590
8
40.5
41.0
42.5
44.5
48.5
44.5
59.5
9
41.0
41 5
430
45 0
49 0
44 5
59 5
10
41.0
42 5
43 5
45 0
49 0
45 0
59 5
11 ....
41.5
42 5
43.5
450
12
41.5
43.0
(a) This quantity of lime was added in addition to the lime already con-
tained in the solution; sufficient lime usually being present in solution that
the addition of 5 c.c. of strong lime water to a titration (with silver nitrate),
for strength of solution, would make no difference in the titration.
(6) See note (a).
(c) Tests No. 5 and No. 6 were on material from near the head of slime
pits, and which therefore contained a larger percentage than usual of fine
sand.
(Each 2 in. of solution equals one ton.)
Before commencing to charge the slime, about 35 tons of solu-
tion from the strong solution sump (usually of a strength between
0.12 and 0.15 of KCN) is pumped into the vat and the attached
centrifugal pump started. From 75 to 100 Ib. of slaked lime is
added and the charging of the slime is commenced. After the
required quantity of slime has been added, a sample of the ma-
terial passing through the pump is taken, filtered and the clear
solution titrated. The necessary quantity of cyanide to bring
the solution up to strength is then added. Experiments have
been made with various strengths of solution in the agitation vats;
• 4
fcl
i:
iU'*fc •
FIG. 82 — General Arrangement of Slime Plant.
FIG. 83 — Three of the Agitation Vats and Tops of two of the Decantation Vats.
CYANIDATION OF AURIFEROUS SILVER ORES 315
the results thus far show the 0.2 per cent, solution to give more
satisfactory results than the use of a weaker solution. The solid
cyanide is placed in perforated buckets or cans and suspended
in the charge. It is found that unless the receptacles containing
the cyanide be frequently agitated about in the charge, the cya-
nide dissolves exceedingly slowly. The less the proportion of
solution to solid matter present, the more noticeable is this ten-
dency of the cyanide to dissolve slowly. It is also noticed that,
the thicker the charge, the slower is the action of the cyanide on
the silver and gold contained in the slime. During agitation it is
best to keep the screen at the end of the suction pipe just as near
the surface of the charge as possible, without allowing the entrance
of air. By so doing, the material passing through the pump
always contains a minimum quantity of solids, and the wear on
the pump is consequently lessened. In addition to this, the
movement or circulation within the charge is then greatest, since
the suction and discharge points are then most separated. It is
quite probable that a considerable portion of the heaviest and
coarsest part of the material treated does not pass through the
pump at all; as, owing to its greater weight, it may never be
raised to the hight of the suction screen. The agitation of the
mass seems to depend chiefly on the fact that the discharge issu-
ing from the drop-pipe tends to keep the point of the conical
bottom free from any settled deposit of slime, and the thickened
material, constituting the lower portion of the charge, keeps
constantly sliding down the inclined sides toward the bottom
point. The product issuing from the discharge-pipe, being drawn
from the surface of the charge, must pass upward through the
entire mass above, before it can again pass through the pump.
The percentage of solid matter contained in the material
passing through the agitation pumps is determined from samples
taken through the bib-nosed petcock tapping the discharge-pipe
a few inches above the pump shell. The pulp passing through
the pumps will carry 25 per cent., by weight, of solids.
A thorough oxygenation of the mass is found to be an essential
feature; it becomes more necessary as the proportion of solid
matter to solution present increases. At the commencement of
the operations, the small air-cock e (Fig. 78) was used to permit
the continuous admittance of air to the suction pipe of the pump.
This practice, however, was soon abandoned, because the agita-
HYDROMETALLURGY OF SILVER
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CYANIDATION OF AURIFEROUS SILVER ORES 317
tion was seriously affected by it. The entrance of air into the
suction pipe had a detrimental influence on the capacity of the
pump, and the effect was found to be injurious to the best agita-
tion. Perhaps the chief trouble was due to the rapid rising to the
surface of the imprisoned air immediately on being expelled from
the discharge-pipe. The air bubbles breaking, on reaching the
surface of the charge, caused a splendid surface movement that
might be easily mistaken for the thorough agitation of the entire
mass without effecting a proper scouring of the bottom point of
the vat. The present practice is to allow the entrance of a smaller
quantity of air into the mass, through the small air-cock I (Fig. 78).
Ordinarily a charge is agitated in the vats from forty to forty-
four hours, after which it is discharged into the corresponding
decantation vat, where it is usually given a two days' treatment.
Should the charge from the agitation vat not fill the decantation
vat, enough precipitated solution is pumped up from the strong
solution sump to fill it; after agitation for half an hour, the charge
is allowed to settle. Should the addition of this extra solution
be unnecessary, the charge is not agitated, but allowed to settle
as long as practicable, the clear supernatant solution being mean-
while decanted off. After the first settling and decantation, the
vat is pumped full of weak, precipitated solution, which is usually
of a strength approximating 0.1 per cent, of KCN per ton, and
the charge is agitated for an hour or two by means of the 3-in.
centrifugal pump connected with the vat, about 25 Ib. of slaked
lime being added during the agitation. The pump is then stopped
and an additional quantity of slaked lime, usually about 10 Ib., is
sprinkled evenly over the top of the charge. After settling a
few hours, the decantation pipe is lowered and the settling and
decanting of clear solution continued as long as practicable. As
many washes and decantations as possible within the time limit
of the treatment are given in this manner. When permissible,
the last wash given is of clear water, though a few of the charges
have to be washed entirely with weak solution.
When treating charges containing the equivalent of 15 tons of
dry slime, usually four settlings and four decantations can be
effected with the forty-eight hours of treatment, each decanta-
tion averaging about 22 tons of solution; hence about 90 tons of
solution is decanted in treating a 15 ton charge, and each decan-
tation removes approximately 58 per cent, of the total solution
318
HYDROMETALLURGY OF SILVER
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^ :5 53 — <
5.9 « cs
w« u-g
o|^H
11 ill
= S »J='S
13 p
fl.i
Jilll-i
Qt ^3 fs^ c3 O rt
$ J^ ^-S E
CYANIDATION OF AURIFEROUS SILVER ORES 319
present. Assuming the wash-agitation to be perfect, the four
decantations should then theoretically contain about 97 per cent,
of the total value dissolved at the time the -washing was com-
menced. The settled pulp is discharged through the bottom
valve and the 4-in. discharge-pipe into the waste launder.
Table III shows the rate of settling per hour, determined at
various times on several different charges.
The pulp, ready for discharging, carries 50 per cent. of moisture,
the contained solution averaging 0.07 per cent, of KCN, and hav-
ing an average value of approximately $0.40 of gold and 1.50 oz.
of silver per ton. These values are higher than would be expected
to remain in the solution after the several decantations and dilu-
tions effected; yet (as has already been recorded by several dif-
ferent parties operating similar slime plants) the solution of value
from the slime does not cease at the completion of the agitation
proper, but continues throughout the washing; the value of the
wash-solution is thus being constantly augmented. This feature,
however, is more noticeable with the silver than with the gold,
and the maximum extraction of the gold is obtained earlier.
For these reasons, the solution contained in the discharged pulp
will always carry more values than it should according to calcula-
tion based solely upon successive dilution and assuming the agi-
tation to be perfect.
A portion of the sample of the pulp ready for discharging
(together with its proper proportion of contained solution) is
dried, the assay results being taken to represent the value of the
discharged slime. Another portion of the pulp is washed and
then assayed. On an average the washed sample will run about
$0.40 of gold and from 1 to 2 oz. of silver per ton lower than the
unwashed sample.
The 3-in centrifugal pumps connected with the decantation vats
are the ordinary pumps commonly used for pumping solutions;
the only alterations being that the bearing nearest the pump
shell is tapped with a ^-in. pipe, which supplies the bearing with
solution under pressure. These pumps run about four hours in
each twenty-four, and have given excellent satisfaction, the only
repair work being an occasional repacking of some of the stuffing-
boxes.
Tables IV and V, giving a somewhat detailed record of the
treatment of one charge, may be taken to represent the usual
320
HYDROMETALLURGY OF SILVER
practice. The usual charge is now but 15 tons of slime (dry
weight), while the proportion, by weight, of solution to slime has
been increased to 2.5 : 1.
Precipitation. — All solution leaving the decantation vats is
passed through the zinc-boxes before being reused. The zinc-
boxes have to be watched closely; owing to the excess of lime
present in the solution, difficulty is experienced in obtaining good
precipitation. Records are kept of the quantity of solution daily
passing through the boxes, together with the assay values of the
solution before and after precipitation. These records show
that, during the last three months, an average of practically 48 tons
of strong and 117 tons of weak solution, or a total of 165 tons,
was passed through the boxes daily; the average assays of the
solution were as given herewith:
TABLE VI
STRONG SOLUTION
WEAK SOLUTION
GOLD
SILVER
. GOLD
SILVER
Entering zinc-boxes
$1.05
0.10
Oz.
2.90
0.40
$0.60
0.10
Oz.
1.70
0.35
Leaving zinc-boxes
The zinc-boxes have a combined total shavings capacity of
approximately 120 cu. ft. ; the rate of flow of the solution through
the boxes during 1904 averaged 1.37 ton per cu. ft. of shavings
per twenty-four hours.
The highest-grade precipitate yet recovered from the slime
plant assayed approximately $6800 gold and 17,300 oz. of silver
per ton.
Tonnage, Percentages, etc. — The normal capacity of the plant,
while treating 15-ton charges, and allowing a two days' treatment
in both agitation and decantation vats, is 30 tons per day. Dur-
ing the last quarter of 1904 approximately 2550 tons of slime (net
dry weight) was treated, and the extraction during this period
(shown by the differences between assays of the charge and of the
residue) was 74.9 per cent, of the gold and 49.2 per cent, of the
silver. During this period the assay value of the slime averaged
$4.35 of gold and 19.25 oz. of silver. During the last two months
(March and April, 1905) 3.56 Ib. of sodium cyanide (equivalent to
CYAN1DATION OF AURIFEROUS SILVER ORES 321
4.40 Ib. of potassium cyanide) was used per ton of slime treated.
The average extraction of silver for the last three months has
been 51 per cent. The consumption of cyanide, zinc and slime
per ton of dried slime treated during this time was: Sodium cy-
anide, 4.42 Ib.; zinc, 0.957 Ib., and lime, 13.95 Ib. The sodium-
cyanide consumption is equivalent to 5.52 Ib. of potassium cy-
anide.
Table VII gives the operating costs per ton of slime treated.
TABLE VII. — SLIME COSTS PER TON
Cyanide (4.42 Ib. @ $0.63) . $2.785
Zinc (0.957 Ib. @ $0.30) 0.287
Lime (13.95 Ib. @ $0.0118) 0.165
Other supplies 0.238
Lubricating 0.033
Labor 0.491
Salaries 0.748
Assay office (labor and supplies) 0.066
Power (ditch maintenance and supplies) 0.621
Miscellaneous (lighting, etc.) 0.002
Management and general expenses 0.179
Total i $5.615
CYANIDING AURIFEROUS SILVER ORES AT SAN SALVADOR, C. A.
The following interesting notes on the working of this process
in San Salvador were communicated to the Engineering and
Mining Journal, June 1, 1905, by Alfred Chiddey:
The process was first applied to tailings from the pan amal-
gamation in which the ore was first roasted with salt in the ordi-
nary way. After all these tailings had been worked, the process
was applied to raw ore, with results so successful that roasting
and amalgamation are now discontinued, and much poorer ore
can be treated. The silver occurs presumably as sulphide and
the ore contains a little copper, which apparently helps the ex-
traction; at any rate, it seems to do no harm.
The mode of working is as follows: The ground ore, after
leaving the arrastras (which are fitted with 30-mesh screens),
is passed over amalgamated plates and allowed to settle in masonry
1 $5.615 Mexican currency during this period was equivalent to $2.66 gold.
The cost of realization on the precipitate is not included in the working cost:
these expenses are high. The average cost of realization on the precipitate
produced in the slime plant, per ton of dry slime, is: Government taxes,
$0.856; treatment charges (including transportation expenses), $1.202;
total, $2.058.
322 HYDROMETALLURGY OF SILVER
tanks, where a rough classification is effected into sand and
slimes. The sand is charged over gratings with 25 Ib. lime per
ton into the percolation vats. A cyanide solution of 0.40 per
cent, strength is introduced, equivalent to nearly one-half the
weight of the sand. After standing twelve hours, the cock is
opened and the vat allowed to drain, the solution passing to the
zinc-boxes. On the following day the surface is leveled off and
raked over. The charge is then allowed to stand for six days
exposed to the air, without adding more solution. On the expira-
tion of six days, a 0.20 per cent, solution is added, and the leach-
ing is continued rapidly without intermission for four days, at
the end of which time a water-wash is employed.
The first wash that comes off, after the vat has been standing
dry for six days, often contains from 12 to 20 oz. silver per ton,
which will run down the second day to 6 oz., on the third day to
3 oz., on the fourth day to 1^ oz.; the water-wash will generally
be from 0.5 to 1 oz. per ton. The sand contains from 13 to 15 oz.
silver and from $3 to $5 gold per ton before treatment. • The
extraction for the past 18 months has averaged from 85 to 90 per
cent, of the silver and 90 to 92 per cent, of the gold.
The consumption of cyanide is a little under two Ib. per ton.
Lately sodium cyanide has been used with good results. It is a
little cheaper, but much more deliquescent than potassium cy-
anide, and during the rainy season a box has to be used up im-
mediately it is opened.
The ore contains argentite, chalcopyrite, sometimes specks
of fahlerz (tetrahedrite), but the amount of copper is small,
generally under one-tenth per cent. The sump solutions always
contain copper, but presumably as sulphocyanate. The precipi-
tation of gold and silver in the zinc-boxes is practically perfect.
The slime is air-dried and mixed with the sand in proportion
of 1 : 2, but the time of treatment is prolonged to 18 days in-
stead of 12, the time required for the sand. The air-dried slime
is previously put through a disintegrator. This method of treat-
ing the slime is only temporary. Although the extraction is
satisfactory the time of treatment is too prolonged, and the
method is available only during the dry season.
The following are results on a few charges of sand:
CYANIDATION OF AURIFEROUS SILVER ORES
323
No. OF
CHARGE
BEFORE LEACHING
TAILINGS AFTER LEACHING
SILVER Oz.
GOLD Oz.
SILVER Oz.
GOLD Oz.
44
20.17
0.37
2.52
0.06
45
15.50
0.30
0.95
0.05
46
14.16
0.30
1.57
0.03
47
16.50
0.36
1.87
0.35(?)
48
13.10
0.30
1.44
0.03
49
16.00
0.30
1.37
0.03
50
14.70
0.30
1.37
0.03
55
15.50
0.26
1.50
0.02
56
13.70
0.23
1.00
0.02
57
13.90
0.21
1.00
0.02
TREATMENT OF ROASTED ORE
As stated above, complex silver ores heavily charged with
metal sulphides are not suitable to be treated raw with a cy-
anide solution on account of the decomposing action of these
sulphides on potassium cyanide. To make this process possible,
such ores have first to be roasted with salt to reduce the con-
sumption of cyanide. The roasting has to be done as carefully
as for the sodium hyposulphite process, in order to secure a mini-
mum loss of silver by volatilization. This is accomplished by
roasting at a very low heat without even raising the temperature
toward the end, so that as few of the metal chlorides are expelled
as possible. These metal chlorides, if expelled by heat, are the
sole cause of the loss of silver by volatilization (silver chloride
as such not being volatile), and it would be folly to employ this
means to remove them from the ore. It is much more rational to
remove them by leaching with water, though the resulting base-
metal solution will have to undergo a treatment similar to that
practised in the lixiviation with sodium hyposulphite, to recover
the silver dissolved therein. Having the ore prepared for ex-
traction as far as that, it seems to be more rational to extract the
silver chloride and the gold subchloride first with sodium hypo-
sulphite, and then, after the latter has been replaced in the ore
by water, to apply the solution of potassium or sodium cyanide
to extract the remainder of the gold and that part of the silver
which was not chloridized. This procedure is advisable and much
improved extraction can be thus expected, for the following
reasons :
It will make the application of the cyanide solution possible
324 HYDROMETALLURGY OF SILVER
to a much larger variety of ores than if the same is applied directly
after roasting and washing of the ore. Not all the metal salts
formed in chloridizing roasting are soluble in water; some will
remain in the ore. like cuprous chloride, lead sulphate and others,
and will act decomposingly on the cyanide solution. Most of
such salts, however, dissolve in a solution of sodium hyposulphite,
and are removed from the ore to a great extent during silver
leaching, so that the cyanide solution will thus be applied to a
much cleaner ore, which may make possible the treatment of
ores which contain a larger percentage of copper, lead and other
metals. The action of the cyanide solution will be more energetic
on those particles which escaped chlorination, and suffer less by
decomposition, if the main bulk of the silver and other metals
has been removed. This is also the case with regard to the gold.
It will be found that, in the case of a silver ore rich enough in
gold to show after roasting with salt a perceptible amount pf
gold when concentrated in a horn spoon, the gold assumes a
much brighter color if this is done with a hyposulphite solution.
In treating rich silver-gold concentrates, as related in Chapter
XXII, I experienced the fact that by first roasting with salt, then
extracting the gold by Plattner's method, and finally the silver
with sodium hyposulphite, only 50 per cent, of the gold could be
extracted, while by leaching the silver first and then applying
the Plattner method, the gold extraction was as high as 95 per
cent. There is no reason to doubt that, from an auriferous silver
ore which was first chloridized, the cyanide solution will extract
the gold more quickly and more completely if the ore be first
leached with sodium hyposulphite.
The increase in cost will be but very slight, as it will not in-
volve any handling of the charge, but will require only a continua-
tion of the leaching with another solvent. The total time may
even prove to be shorter than if only a cyanide solution was used.
It seems that very high extractions of silver and gold may
be obtained by a combination of the two processes, and thorough
experimenting on that line will in all probability be well rewarded.
This refers, of course, only to higher grade ores that can stand
the cost of roasting, and the nature of which resists a high chlori-
nation.
John F. Allan, City of Mexico, in a very interesting paper on
"Cyanide Treatment of Silver Ores in Mexico," read before the
CYANIDATION OF AURIFEROUS SILVER ORES 325
Atlantic City meeting, February, 1904, of the American Institute
of Mining Engineers, gives two examples of cyaniding silver ores
containing gold which were roasted with salt before leaching:
"Leaching: Example No. 1. — The plant in question has a
capacity of 700 tons per month. The chloridized pulp is charged
into filters 18 ft. in diameter and 4 ft. deep, and receives as pre-
liminary treatment four hours' soaking with water and seven hours'
percolation. This water-wash is passed through special zinc-
boxes, where an impure precipitate containing gold and silver
is formed, and a consumption of about 0.4 Ib. of zinc per ton
takes place. When the salt and impurities have been removed
by the wash-water, the charge receives fifteen hours' soaking and
ninety hours' percolation with cyanide solution of 0.3 percent,
or 6 Ib. to the ton of water, the ore receiving an equivalent of 1.5
of solution to 1 of ore, or 0.45 per cent. To displace the strong
solution, twelve hours of percolation with a 0.05 per cent, weak
solution is given, in the proportion of 1 of water to 2 of ore, and
a final thirty hours' percolation with water-wash. The con-
sumption of cyanide is 2.15 Ib. per ton, and zinc 1.08 Ib.
"The extractions are: Gold, 76; silver, 85.25; total value, 81.77
per cent. It will be observed that the gold extraction is not as
good as the silver, a fact often noted in chloridized ores. The
following are the costs, which can be considered as typical in small
plants, although in some instances they have been reduced:
Superintendence, $0.62; labor, $0.96; cyanide, $1.71; zinc, $0.29;
laboratory, $0.10; various stores, $0.50; total, $4.18 Mex., or,
at 50c. U. S. per $1 Mex., $2.09 U. S. per ton."
"Leaching: Example No. 2. — The plant is capable of treat-
ing 1800 tons per month. The treatment, with small variations,
is practically the same as in Example No. 1. The consumption
of cyanide is 1.73 Ib. and of zinc 1.27 Ib. per ton, the high consump-
tion of zinc being due to the ore averaging over 30 oz. of silver
per ton. The extractions are: Gold, 82.30; silver, 77.32; total
value, 79 per cent. The costs are: Superintendence, $0.33;
labor, $0.31; cyanide, $0.99; zinc, $0.33; laboratory, $0.10; vari-
ous stores, $0.22; total, $2.28 Mex., or, at 50c. U. S. per $1 Mex.,
$1.14 U. S. per ton."
It is to be regretted that no analysis or description of the ore
is given. The fact that in Example No. 1 the gold extraction
is only 76, and the silver 85.25 per cent., and Mr. Allan's remark
326 HYDROMETALLURGY OF SILVER
that the fact of a lower gold extraction is often noted in chlori-
dized ores, verify my experience in treating auriferous silver
ores. If to the roasted ore of Example No. 1 a solution of sodium
hyposulphite had been applied before the cyanide solution, the
extraction of both metals would have been higher, and very
likely that of gold better than the extraction of the silver.
TESTING THE CYANIDE SOLUTION FOR GOLD AND SILVER
In order to conduct the cyanide process intelligently frequent
tests of the solutions for gold and silver are necessary.
Arents' Test. l — This test is based upon the fact that metallic
copper will precipitate gold and silver upon its surface from
acid solution. Of course the fact is not new, but its application
is probably so. Arents has used the method with success; it
recommends itself by the rapidity and ease with which it may be
carried on.
An auriferous cyanide solution, if made acid with sulphuric
acid and boiled with finely divided, pulverulent, metallic copper,
will, within a short time, deposit its gold content on the copper.
Any silver in the solution is also precipitated. If this mixture
is now filtered, the filter and contents may at once be subjected
to a crucible assay treatment, and its lead button cupeled and
determinated.
If, instead of taking cement copper, or any metallic copper
powder, a solution of bluestone is used after acidification, and a
few small pieces of sheet aluminum are added, and the solution
boiled until all the copper has come down, the result as to the
precipitation of gold and silver is the same. This modification
takes more time and attention in boiling. If aluminum has
been used, it should go into the crucible with the filter and its
contents. Commercial cement copper is particularly fitted for
this test, because the acid, in taking up any basic iron or copper
salts of the cement copper, renders the copper as finely divided
as it is customary to obtain in the sluice-boxes of copper leachers.
The finer and the more pulverulent the copper is, the greater is its
surface and the more energetic the precipitation, thus permitting
a minimum amount of copper to be used.
In applying the method, Arents uses, as a rule, 250 c.c. of the
1 From a paper by Albert Arents read before the Albany meeting, Feb.
1903, of the American Institute of Mining Engineers.
CYANIDATION OF AURIFEROUS SILVER ORES 327
solution to be tested and a few c.c. of sulphuric acid, agitates for
several seconds, and then adds not less (although not much more)
than one gram of cement copper. Now follows heating to boil-
ing. This is kept up for about 10 minutes, so that the rising
steam-bubbles keep the mixture well agitated. The mixture is
then filtered through a 7-in. diameter gray filter-paper. No
washing is done. As soon as the filtering is finished, one-third of
a crucible charge of flux is added to the filter containing all the
sediment of the mixture. Some of the moisture is rapidly absorbed
by the flux, which permits the folding of the filter's rim upon the
charge and its subsequent removal without loss or tearing. One-
third of a crucible charge of flux having previously been placed
upon the bottom of the crucible which is to be used for melting,
the filter is transferred to the crucible, well tucked down, and the
last third of the crucible charge is placed on top of the filter in the
crucible. It is then ready for the furnace. The filter itself fur-
nishes the reducing agent for the assay. Arents uses 30 grams
litharge and the usual amount of borax and soda, employing a
No. F crucible for melting. About 20 grams of lead are obtained.
The lead button comes out bright and clean, and upon cupeling
furnishes a bead free from copper.
Possibly this method of testing for gold and silver may be
used upon other solutions than cyanide; also, for solutions from
testing metallic copper for precious metals, when the solutions do
not contain nitric acid in any form.
Alfred Chiddey's Test. — Four assay tons of solution are taken
in a porcelain dish, and 20 c.c. of a 10 per cent, solution of lead
acetate added; then 4 grams of zinc shaving and afterward 20 c.c.
of hydrochloric acid. When the action has nearly ceased, the
contents of the dish are boiled for a minute or two and filtered.
The precipitate is well washed with water, moistened with alco-
hol, dried, wrapped with the filter in lead foil, and cupeled.
The testing of the strength of a solution in potassium cyanide
is usually done by titrating the same with a standard solution
of silver nitrate. This standard solution is prepared by dissolv-
ing 13.04 grams of c.p. silver nitrate in one liter of distilled water.
Every 0.1 c.c. of such a solution added to 10 c.c. cyanide solu-
tion is equivalent to 0.01 per cent, of potassium cyanide. The
operations are as follows:
A Mohr's burette, graduated to 0.1 c.c., is filled with the
328 HYDROMETALLURGY OF SILVER
standard silver solution. Ten c.c. of the cyanide solution which
is to be tested is taken up with a pipette and emptied into a small
flask which is brought under the burette. Then drop by drop the
silver solution is added. Each drop produces a whitish precipi-
tate, which, however, disappears again when the flask is well
shaken. This is continued until the last drop produces a cloudi-
ness or turpidity, which remains even after a vigorous shaking of
the flask.
The scale on the burette is now read. Each tenth of a c.c.
silver solution used indicates 0.01 per cent, of potassium cyanide
contained in the solution.
INDEX
Agitating tanks 263
Agitation vats 304, 305
Agitators, in precipitation vats. 185
Air blow-off drum 200
compressed, as agitator in
precipitation vats 185
effect of, in roasting calca-
reous ore 147
with highly sulphureted
ore 134
provision for, in Bruckner
furnace 69
required in chloridizing
roasting 4, 56
Allan, John F '.'. .324, 325
Alumina, causes loss of silver in
roasting 8
Aluminum chloride, formed in
roasting 8
Amalgamation Pref . iii
barrel 37
pan 37
roasting for .... 37
volatile chlorides
undesirable in 20
American underfed stoker 272
Analysis of ores of Anglo-Mex-
ican Mining Co.,
Yedras, Mexico 127
of precipitate of ores
of A vino, Mexico ... 195
of San Francisco del
Oroore 100,118
Anglo-Mexican Mining Co., Ye-
dras, Sinaloa, Mexico, analysis
of ores of 127
roasting ores of .... 50
Anhydrous sulphuric acid, for-
mation of 4
Antimonial fahlerz, effect of
steam on ores con-
taining 33
galena, effect of
steam on ores con-
taining 33
minerals, effect of, on
time of lixiviation 180
silver minerals, be-
havior of, in roast-
ing 9
Antimony antimonate, formed in
roasting 7
expelled by steam in
roasting 31
sulphide, behavior of,
in roasting 7
trichloride, action of,
in roasting 7
Appearance of ore in steps of
roasting 5
Arch, construction of in long
reverberatory furnace 55
Arents, Albert 326, 327
Arents' test for gold and silver. 326
Argentiferous black copper, ex-
traction of silver
from 278
copper matte, used
for sulphating
roasting 94
zinc-lead ore, chlo-
ridizing of .... 99
Argentite, behavior of , in roasting 9
Arsenate of silver, formed in
roasting. . . 7
soluble in so-
dium hypo-
sulphite . . 7
329
330
INDEX
Arsenic, action of, in roasting. . 7
chloride, formed in
roasting 7
expelled by steam in
roasting 31
Arsenical minerals, effect of on
time of lixiviation . . 180
ores, at Yedras, Mex-
ico, experi-
ments with . . 127
long furnace for. 50
pyrites, behavior of, in
roasting 7
silver minerals, behav-
ior of, in roasting. . 9
Arsenious oxide, formed in roast-
ing 7
Assay values of raw and roasted
ores of Cusihuiriachic Silver
Mining Company 24
Augustin process 155, 256
Auric chloride, action of, in
roasting 35
Auriferous black copper, extrac-
tion of silver
from 278
silver ores, at Pal-
marejo, Mexico 288
Kiss process for. . 254
roasting of 34
treatment, Pref . iv, 256
Aurous chloride, action of, in
roasting 35
Avino, Durango, Mexico, analy-
sis of precipitate of ores of . . 195
Bag system of collecting dust . . 91
Ball-mill 143
Balling of the ore 142
Barrel amalgamation 37
Base-metal chlorides, action of,
in lixiviation
troughs 222
desilverizing by
water 170
expulsion of, in
amalgamation 38, 39
Base-metal solubility of silver
chloride in ... 161
see also Metal
chlorides
leaching 157
at Sombrerete,
Mexico 166
cause of chlorina-
tion 115
cupric chloride
. used during. .167,
173
description of . . 156
of silver ores rich
in gold 284
tanks for 229
time required for 244
solutions, loss of
silver in . . 162, 246
precipitation of
silver from ... 164
Battery, stamp, adding salt in. . 140
effect of crush-
ing ore in ... 11
Bins, construction of, for heap-
roasting 31
Bisulphide of calcium, formed in
calcium sulphide tank 190
Bituminous coal, used in chlori-
dizing roasting 42
used in sulphating roasting . . 95
Black copper, extraction of silver
from 258,278
Blake rock-crusher 289
Blue vitriol, formed in extrac-
tion with sulphuric
acid 259,275
in base-metal leach-
ing at Sombrerete 169
Bosque mill at Parral, Mexico 101, 243
Bottom of long reverberatory
furnace, construction of 52
Brown furnace 71
Bruckner 63
Bruckner cylinders, tests for
loss of weight in . 23
furnace 144, 147
INDEX
331
Bruckner furnace, action of salt in 18
advantages of . . 66
advantages over
reverberatory
furnace 30
behavior of ore in 140
decomposition of
soluble silver in 146
dust formed in . . 87
effect of adding
salt in 135
extra handling in 143
Hofmann im-
proved 67
loss of silver in,
less than in re-
verberatory
furnace 20
not suited for
roasting zinc
lead ores 126
ores suitable for. . 67
roasting calcare-
ous ore in ... 128
roasting sulphu-
reted ores in . . 42
self-roasting pro-
cess in 152
sulphureted ores
favorable to
roasting in ... 150
used at Yedras,
Mexico 17
used in chloridiz-
ing self-roasting 26
used in experi-
ments with ore
from Silver King
mine 34
revolving furnace . 63
speed of revolu-
tion 66
roaster 63
Burlap, used for filter in leaching
tanks 183
Cadmium, in zinc blende ore ... 99
Calcareous arsenical silver ore at
Yedras, Mexico, be-
havior of, in
roasting 17
experiments with 127
loss of silver in
roasting 21
ores, chloridizing . . . 127
long furnace for
roasting 50
Calcined copperas, added in oxi-
dizing roasting 18
Calcium chloride formed in roast-
ing 8
hyposulphite, action of,
in silver leaching . 179
used in Kiss process . 254
polysulphide, prepara-
tion of 186
sulphate, cause of ball-
ing 143
. sulphide, as precipitant 164,
179, 183, 198
as test for sil-
ver 178
boiler and pres-
sure tank for 188
Calcspar, in gangue of ore from
San Francisco del Oro mine. 100
Capacity of long reverberatory
furnace 60
Carbonate of lime, as material
for cupel . . 210
behavior of, in
roasting .... 8
effect of, in
roasting ore 131
used to decom-
pose base
chlorides . . 39
Caustic lime, action of, in roast-
ing calcareous ores. .8, 134
potash, manufacture of 205
soda, effect of, on silver
chloride 204
Cement copper, as precipitant
for silver 165, 167, 257
332
INDEX
PAGE
Centrifugal pumps 304, 319
Challenge feeders 289
used in the Ropp
furnace 77
Charge hopper, construction of,
in long reverberatory furnace 55
Charging a long reverberatory
furnace 57
Chemical loss in weight during
roasting 20
Chiddey, Alfred 321, 327
Chiddey's test for gold and silver 327
Chloride of copper, used to cor-
rect bad roasting 104
Chloridizing heap-roasting 27
advantages of . . 31
of calcareous ores. 127
period of roasting. 4
roasting, definition
of Pref . iii
in relation to
amalgamation
Pref. iii
modification of,
for lixivia-
tion Pref. iv
of argentiferous
zinc-lead ore,
conclusions of
experiments . Ill
experiments .... 101
theory of 3
with steam .... 31
see also Roasting
self-roasting 26
in Bruckner revolving
furnace 65
Chlorination, after ore has left
furnace 115
aided by base-
metal leaching 115
effect of salt on . . 16
gold 36
lowered by excess
of salt 107
of silver, methods
of effecting 3
Chlorination on cooling floor . . 27
Chlorine, formed in roasting 3,
6, 8, 29, 30
Clay, es material for cupel 210
behavior of, in roasting . . 8
Clean-up box 301
of zinc-boxes in cyani-
dation 300
Cleaning a Bruckner furnace. . . 71
Coarse crushing, advantages and
disadvantages of 12
Compressed air, as agitator in
precipitation
vats 185
in pressure
tanks 199
Consolidated Kansas City Smelt-
ing and Refining Company,
Argentine, Kansas 215, 261
Continually discharging furnaces,
fuel needed in 42
Continuous feeding mechanical
roasting furnaces 71
Cooling floor, Chlorination on ... 27
Copper, as precipitant for silver 165
effect of, in ores 15
in silver leach-
ing 178
on time of
lixiviation . 180
matte, extraction with
sulphuric acid
from 260
in Augustin pro-
cess 256
suited for Zier-
vogel process 281
pyrites, behavior of, in
roasting 6, 9
in ore from San
Francisco del
Oro mine ... 100
removing from precipi-
tate before cupellation 195
sulphides, advantage of,
inchloridiz-
ing ores ... 4
INDEX
333
Copper sulphides, necessary in
ores for sul-
phating
roasting. . . 94
Cost of cyaniding auriferous sil-
ver ore 303
of roasting in the modified
Howell furnace 119
Crushing in stamp batteries ... 11
ore from San Francisco
del Oro mine, exper-
ments in 11
through rolls 11
Cupeling furnace, dust-collector
for 213
refining precipitate
in 179,209
Cupellation, refining precipitate
by 195
Cupric chloride, action of, in
chloridizing silver . . 4
as precipitant for silver 168
formed in roasting .... 6
in base-metal leach-
ing 167,173
in water an aid to
chlorination 115
preparation of 173
treating ore with .... 125
used in base-metal
leaching 173
used in foul hypo solu-
tion 181
oxide, action of, in roast-
ing 6
sulphate, formed in roast-
ing 6
formed in sulphating
roasting 94
product of extraction
with sulphuric acid . 258
Cuprous chloride, effect of, in
leaching . . 177
formed in
roasting. . . 6
in base-metal
leaching . . 168
Cuprous oxide, action of, in
roasting 6
Cusihuiriachic, Chihuahua, Mex-
ico 254
lixiviation at 155, 240
loss of silver in leaching
ore from 163
time required for lixivia-
tion at 180, 182
working ore from 14
Cusihuiriachic Silver Mining
Company, Chihuahua, Mexico 24,
80, 221
Cyanidation of sand 292
Cyanide leaching plant 293
leaching vats 292
process for silver ores
Pref. iv.
Cyaniding auriferous silver ores,
at Palmarejo, Mexico . . 288
at San Salvador, C. A. . 321
consumption of cyanide,
zinc, and lime 302
description of slime
plant 305
in Mexico 325
precipitation 320
precipitation of silver
and gold 299
sizing test on slime. . . . 305
time required 317
tonnage and extraction . 302
tonnage, percentages,
etc 320
treatment of roasted ore 323
treatment of slime. . . . 304
working costs 303
Dead roast 6, 18, 259
Decantation vats 304, 311
Decrepitation, caused by salt in
a Bruckner furnace 18
Del Oro ore, see San Francisco
del Oro ore
Desilverization of waste liquor. 165
Desilverizing base-metal chlo-
rides with water . .170
334
INDEX
Distillation of sulphur from pre-
cipitate 207
Distributing trough for milk of
lime 188
Don Enrique Mining Company,
Cusihuiriachic, Chihuahua,
Mexico 240
Drum, air blow-off 200
Drying and roasting furnace for
silver precipitate 208
Dust in White-Howell furnace 80
collecting methods, bag
system 91
Hofmann's flue-dust col-
lector 88,214
Dusting of ore 27
English cupeling furnace 209
Experiments in roasting ore
from San Francisco del Oro
mine 101
Extraction, of the gold from rich
silver-gold ores . . 284
with sulphuric acid. 258
Fahlerz, behavior of, in roasting 6
Ferric arsenate, formed in roast-
ing 7
chloride, action of, in
roasting 5,6,7
oxide, formed in roasting 6, 7
sulphate, formed in roast-
ing 6, 7
Ferrous chloride, action of, in
roasting 5,6,7
sulphate, added in oxi-
dizing roast-
ing 18
formed in sul-
phating roast-
ing 94
Filter bottom of cyanide leach-
ing vats 292
construction of 159
press 267
Johnson 199
Filtering precipitate 204
Filtering quality of ore improved
by lead sulphide 14
Filters for leaching tanks 182
Filtration, effect of coarse and
fine crushing on . . 12
effect of, on time of
lixiviation 181
Fine crushing, effect of, on fil-
tration 12
Flint, Idaho, treatment of ores
from « 40
Flue-dust collecting methods . . 87
collector, Hofmann's 88,
214
-hole, construction of, in
long reverberatory fur-
nace 55
Free filtration, effect of, on time
of lixiviation 180
percolation, conditions
which aid, in roasting. 12
effect of coarse and fine
crushing on 12
Freiberg, Saxony, extraction
with sulphuric acid at 258
Fuel, consumption of, in roast-
ing 32,42,119
Furnace, see Bruckner, Howell,
Howell-White, Long reverber-
atory, McDougal, Mechanical
roasting, O'Harra, Pearce, Re-
verberatory, Ropp, Stetefeldt,
Two-story reverberatory
Galena, behavior of, in roasting 6, 9
effect of, on time of
lixiviation 180
effect of salt on, in
roasting 16
steam in roasting ores
containing 32
in ore from San Fran-
cisco del Oro mine . . 99
not desirable in a charge 6
not favorable in roast-
ing in Stetefeldt fur-
nace.. 86
INDEX
335
Gangue, effect of, on free perco-
lation 12
minerals containing
alumina, behavior of
in roasting 8
of ore from San Fran-
cisco del Oro mine . 100
Gas, producer, as fuel for roast-
ing 42
Gay-Lussac tower 91
German cupeling furnace 209
Gerstenhofer pyrites roaster . . 83
Glauberite, cause of balling .... 143
Globules of ore, effect of steam
on, in roasting 32
Gold chlorides, action of, in
roasting 35
chlorination 36
cyanidation of silver ores
rich in 287
extraction of, from rich
silver-gold ores 284
precipitation of, in cya-
niding 299
roasting of silver ores con-
taining 35
testing cyanide solution
for 326
treatment of silver ores
rich in 283
Gravel, as filter in silver leach-
ing 182
Gray copper ore, behavior of, in
roasting 6, 9
Grinding machine 238
Grizzlies 289
Gypsum, removal of, in Zier-
vogel process 282
Hand-worked furnaces 45
reverberatory fur-
naces vs. me-
chanical roast-
ing 62
Hauch 155
Heap-roasting 27
advantages of . . 31
Heat, amount of, required in long
reverberatory furnace. . 59
excess of, cause of loss of
silver 20
Hidalgo Mining Company, Parral,
Mexico 170, 171
Hofmann improved Bruckner
furnace 67
Hofmann 's flue-dust collector 88, 214
methods of extract-
ing silver with sul-
phuric acid 259
modified Howell fur-
nace 81
Homestake mortar 290
Howell furnace, fuel needed in 42
Hofmann 's modified . 81
modified 243
cost of roasting in . . 119
roasting San Fran-
cisco del Oro ore
in 113
not successful in roast-
ing San Francisco
del Oro ore 112
preferable to long re-
verberatory 52
provision for air in . . 128
roasting zinc-lead ores
in 112
steam used in 110
tests made in, at Chi-
huahua, Mexico ... 24
used after roasting in
Stetefeldt furnace. 109
used at Chihuahua,
Mexico . 221
Howell-White furnace, descrip-
tion of 77
dust-collector in ...88,91
dust formed in 87
good fuel-economizer 44
in Bosque mill at
Parral 101
remedy for dust in . . 80
see also White-Howell
furnace
336
INDEX
Hydrochloric acid formed by
steam in
Howell
furnace . 1 10
in roasting 31
formed in roast-
ing 3,4,6
used in test for
silver 183
Iodine solution, test for silver. . 178
Iron chlorides the principal chlo-
ridizers 5
effect of, in ores 15
pyrites, advantage of, in
chloridizing ores 4
behavior of, in
roasting 5,9
in ore from San
Francisco del
Oro mine 100
sulphate, action of, in roast-
ing 17
sulphide, action of, in roast-
ing 6
advantageous in sulphat-
ing roasting 94
contained in zinc blende 7
sulphides needed in chlori-
dizing silver 5
needed in sulphating
roasting ores 94
Joachimsthal, Bohemia, lixivia-
tion at 155
Johnson filter press 199
Kiss 179, 254
Kiss process 254
Krupp ball-mill 261, 263
Kustel, G 39, 40, 41, 71, 86
La Baranca, Sonora, Mexico,
gravel used for filter at 182
La Dura, Sonora, Mexico, lixivi-
ation at 155
Labor required on long reverber-
atory furnace 59
Las Bronzas, Mexico, lixiviation
at 155
Leach-troughs 215
Leaching auriferous silver ores . 325
base-metal 157
at Sombrerete 166
description of 156
silver 174
description of 156
method of 177
tanks, construction of 157
filters for 182
to remove base-metal
chlorides 21,40
Lead bath, refining precipitate on 209
carbonate, by-product in
silver leaching 178
chloride, action of in amal-
gamation .... 41
formed in roast-
ing 6
effect of, in ores 15
in silver leaching 178
lining not suited for lixivi-
ation troughs 216
oxide, formed in roasting. 6
silicate formed in roasting 32
sulphate, action of in amal-
gamation 41
effect of, in leaching. . 177
formed in roasting .... 6
sulphide, an effective agent
in filtering ore 14
see also Galena
Lexington mine, experiments
with ore from 28
Lime beneficial to roasting ... 8
milk of, as precipitant for
silver 164
used in preparing cal-
cium polysulphide . . . 186
rock, see Carbonate of lime
Limestone, cause of balling .... 143
gangue, effect of, in
roasting 10
Litharge, used in refining silver
precipitate 208
INDEX
337
Lixiviation, Augustin process . 256
effect of antimonial
minerals on
time of 180
arsenical minerals
on time of. ... 180
Kiss process 254
process, reaction of
calcium sul-
phide and
sodium sul-
phate 198
tests for silver in 178
treatment of pre-
cipitate in .... 198
used on calcare-
ous ores 127
volatile chlorides
not objection-
able in 21
Russell process ... 251
tank 219
time required for. . 180
trough 162, 219
troughs 224
arrangement and
operations of. . 229
construction of. . 215
with sodium hypo-
sulphite 155
description of
process 156
first introduced
Pref. iv
Long, J. T 171
Long reverberatory furnace ... 47
capacity of 60
charging 57
construction of 52
heat required 59
labor required 59
not useful for roasting
ores low in sulphur 52
two-story 60
used on calcareous ores
at Yedras, Mexico.. 128
Loss of silver by volatilization . 117
Loss of silver, determination of
amount 163
in base-metal solutions 162,
246
in Stetefeldt furnace . . 85
in sulphating roasting 97
Lucky Tiger mine, Sonora, Mex-
ico, experiments with ore from 36
Lump-grinding machine 238
Lye, manufacture of, from
wood ashes 205
McDougal furnace, fuel needed in 42
Mansfeld, Germany, process of
sulphating roast-
ing at 95
Ziervogel process at 94,
281
Mechanical roasting furnace ... 62
fed by charges ... 63
vs. reverberatory
hand worked . . 62
with continuous
feeding 71
Hofmann's im-
proved How-
ell furnace . . 81
Howell-White
furnace 77
O 'Harra furnace 71
Ropp furnace. . 74
Stetefeldt fur-
nace 82
Mercury, action of base-metal '
chlorides on 40
effect of amalgamation
on 37
Metal chlorides, formation of
volatile, the cause of
loss of silver 30
formed in roasting .... 3, 4
reaction for formation of 4
silver chloride soluble in 40
volatile, action of, in
chloridizing silver ... 4
see also Base-metal chlo-
rides
338
INDEX
Metal oxides, formed in roast-
ing 8
silicates, action of steam
on, in roasting 31
subchlorides formed in
heap-roasting 30
sulphates, formed in
roasting 3
sulphides, effect of, on free
percolation 12
Mexican Santa Barbara Mining
Company 61
Milk of lime, as precipitant of
silver 164
used in preparing
calcium polysul-
phide 186
Modified Ho well furnace, com-
pared with reverber-
atory 126
cost of roasting in. . 119
results of experi-
ments with. . .116, 120
roasting San Fran-
cisco del Oro ore
in 113,243
Mohr's burette 327
Monitor, California, lixiviation at 155
Muffle, experiments in 143, 196
loss of silver in samples
roasted in 20
Native silver, in ore from San
Francisco del Oro mine 100
Nitric acid, used in test for silver 183
North Mexican Silver Mining
Company, Mexico 69, 240
O'Harra furnace 51, 71
Oker, Germany, extraction of
silver from copper matte from 278
Oker process 278
Ontario, Utah, roasting ores from 41
Ores, silver, classification of in
relation to roasting ... 9
suitable for chloridizing
roasting 3
Oxide of antimony, formed in
roasting 7
Oxidizing period of roasting /. . 4
roasting, experiments
with two-story re-
verberatory fur-
nace 122
necessary for cer-
tain ores 16
Oxnam, T. H 288
Oxygenation of gold ores 296
Palmarejo and Mexican Gold
Fields, Ltd., Chi-
nipas 288
Chihuahua, Mexico . 288
ores 296
Pan amalgamation 37
evaporator 270
Parral, Chihuahua, Mexico, 11, 170,
174
Patio 291
Pearce turret furnace ... 71, 261, 263
Pelton wheels 289, 304
Percy 155
Physical changes in roasting ore 13
Plattner 85
Plattner's method 283, 285, 324
Plomosas, Mexico 41
Plumbiferous silver ores, lixivia-
ation of 220
roasting 41
Porphyry, behavior of, in roasting 8
Potassium cyanide 287
solution used to ex-
tract silver 104
Precipitant, adding, in precipi-
tation of silver 191
Precipitate, pressure tanks for
treatment of. ... 205
refining by cupella-
tion 195
removal of sulphur
from 203
silver, drying and
roasting furnace
for . . . 208
INDEX
339
Precipitate, silver, refining the 208
treatment of.. .194,198
Precipitation 320
of silver 185
adding the pre-
cipitant 191
and gold 299
chloride by di-
lution with
.water 170
from base-metal
solutions .... 164
preparing cal-
cium hypo-
sulphide for . 186
process, descrip-
tion of 156
tanks 156
vats 185, 233
Pressure tanks 199, 200, 205, 265
Producer gas, as fuel for roasting 42
Purifying tower . 267
Quartz, a desirable gangue in
roasting 8
behavior of, in roasting 8
in gangue of ore from
San Francisco del Oro
mine 100
Refining the silver precipitate. . 208
Reverberatory furnace .... 143, 254,
263, 278
compared with Bruckner 30
description of 45
dust formed in 87
experiments with in
heap-roasting 29
fuel needed in, for roast-
ing ores poor in sul-
phur 43
fumes not visible in
roasting in 30
hand-worked, vs. me-
chanical roasting. ... 62
long, see Long reverber-
atory furnace
Reverberatory furnace, loss of
silver in, greater than
in Bruckner furnace . 20
roasting calcareous ore
in 144
roasting plumbiferous
silver ores in 41
rules for temperature
and draft in 149
single-hearth 45
two-story long 60
two-story single-hearth 46
used at Sombrerete. ... 31
used to remove sulphur
from precipitate .... 207
Rising Star mine, Flint, Idaho,
ore from 72
Roasting, appearance of ore in
steps of 5
auriferous silver ore . 323
chloridizing, see Chlo-
ridizing roasting
heap- 27,31
self- 26,65
for amalgamation ... 37
in the Bruckner fur-
nace 128
methods of 26
silver, ores containing
gold 35, 285
steps in process of . . 4
sulphating 94
Rolls 11,143
Roof of long reverberatory fur-
nace, hight of 53
Ropp, Alfred von der 75
Ropp furnace 51, 74
Russell 163, 166, 178, 253
Russell method of precipitating
lead 179
Russell process 251
Russell's extra solution, used to
extract silver 104, 252, 254
Salt, action of, in roasting 4
adding during crushing pro-
19
340
INDEX
Salt, adding in battery 140
addition of, in amalgama-
tion 38
best time to add, in roasting 16
coarse vs. pulverized in
roasting 18
effect of adding in roasting 9
excess of, on loss
of silver 164
in roasting ore
from San Fran-
cisco del Oro
mine 112
in roasting zinc-
lead ore 102
on balling of ore. 142
on calcareous ores 135
percentage of, needed in
roasting 15
pulverized vs. coarse in
roasting 18
San Francisco del Oro ore 17
analysis of raw 100
analysis of roasted 118
chlorination after ore has
left furnace 115
conclusions of experi-
ments in chloridizing 111
cupric chloride used in
leaching 174
effect of salt on 103
experiments in chlori-
dizing 31,99
crushing 11
heap-roasting 28
trough lixiviation . . . 243
furnaces used in roasting 61
loss of silver in leaching 162
in roasting 117
methods of treating .... 252
results of roasting in
modified Howell fur-
nace 115
salt required in roasting. 16
stock solution used on . . 175
time required to work by
lixiviation ... . 248
San Francisco del Oro ore, using
Howell furnace on . . 81
wood consumed in roast-
ing 119
San Marcial, Mexico 155
San Salvador, Central America. 321
Sand, as filter in silver leaching 182
cyanidation of 292
-retaining tank 290
Santa Barbara 11
Schemnitz, Hungary, Ziervogel's
method tried at 95
Self-roasting calcareous ores ... 140
chloridizing 26, 65
Settling-tank for sluicing 225
Silver chloride decomposed by
caustic lime 8
obtained by chloridizing
roasting. / 3
precipitating with water 170
soluble in solution of
metal chlorides 40
solubility of 161
copper glance, behavior of,
in roasting 9
extraction of, by the Zier-
vogel process 281
from black copper .... 278
from copper matte .... 258
with sulphuric acid . . . 258
Hofmann's method of ex-
traction with sulphuric
acid 259
leaching, calcium hyposul-
phite, action of 179
calcium sulphide
as precipitant
in 179
description of . . 156
effect of lead and
copper 178
end of 183
filters for leach-
ing tanks .... 182
in lixiviation with
soduim hypo-
sulphite 174
INDEX
341
Silver leaching in ores rich in
gold 284
in trough lixivia-
tion 246
method of 177
regeneration of
hypo solution 181
tanks for 229
testing for silver
in 178
time required 180,248
loss of, by volatilization 20, 117
determination of
amount 163
in base-metal solu-
tion 246
in roasting, meth-
ods of ascertain-
ing 22
in Stetefeldt fur-
nace 85
in sulphating roast-
ing 97
reduced by steam 33
methods of chlorination of 3
native, in ore from San
Francisco del Oro mine 100
ores, classification of in
relation to roasting 9
rich in gold, cyani-
dation of 287
treatment of 283
precipitate, fineness of . . 248
treatment of . . 194
precipitation of 185
from base-metal solu-
tions 164
in cyaniding 299
prevention of loss of .... 162
recovery of from waste
liquor 165
testing cyanide solution
for 326
Silver King mine, Arizona, Bruck-
ner furnace mod-
ified to roast
ores from 67
Silver King mine, desilverizing
base-metal chlo-
rides at 170
experiments with
ore from ....'.. 33
leaching with cu-
pric chloride at 174
lixiviation at .... 155
regenerating hypo
solution at .... 181
time required for
lixiviation at . . 180
Single-hearth reverberatory fur-
nace 45
two-story 46
Sizing-test on slime 305
Slate, behavior of, in roasting. . 8
Slime, formed in crushing 12
method of treatment ... 313
pits 291
plant, description of ... 305
settling rate per hour . . 314
sizing test on 305
treatment of, in cyanid-
ing auriferous silver
ore 304
Sluice-tanks 225
Sluicing 225
Sodium carbonate, precipitate
for lead 178,179
chloride, addition of, to
ore 3
formed in roasting . . 4
method of decompo-
sition of, in roast-
ing 3
cyanide 287
hyposulphite, a solution
for arsenate of silver 7
action of, in chlo-
ridizing zinc-lead
ores 104
best strength of solu-
tion 177
handling 201
in silver leaching ... 174
lixiviation with ... . 155
342
INDEX
Sodium hyposulphite, regenera-
tion of, when foul. 181
used in extracting gold
from silver ore .... 35
used in leaching zinc-
lead ore 106
used on calcareous
ores 127
silicate, formed in roast-
ing 8
sulphate, accumulation
of, in leaching solu-
tion 197
cause of balling 143
formed in roasting . . 4
sulphide, as precipitant 164,
171
as test for silver 178
Solubility of silver chloride .... 161
Solution tanks 312
Sombrerete, Zacatecas, Mexico 254
base-metal leaching at ... 166
experiments at 30
experiments in heap-roast-
ing ore from 28
loss of silver in leaching
ore from 163
ore from 11, 17
reverberatory furnaces
used at 31
straw for filter used at . . 182
two-story long furnaces at 61
Stamp battery, adding salt in. 140
effect of, in
crushing ore 11
Starch paper, test for silver. . . 178
Steam, applied in Howell furnace 110
purifying tower . 267
roasting by Von
Patera 155
chloridizing roasting with 31
consumption of fuel in
use of, in roasting ... 32
effect of, in roasting .... 32
in base-metal leaching at
Sombrerete 169
loss of silver reduced by 33
Steam pump, used instead of
pressure tanks 199
used in desilverizing waste
liquor 165
used in roasting to aid
the extraction of silver 1 3
used in extraction with
sulphuric acid 259
Stetefeldt, C. A., 27, 41, 83, 128, 166,
208
Stetefeldt furnace, capacity of. 85
conclusions of ex-
periments on zinc-
lead ores in .... Ill
description of .... 82
draft required in.. 109
dust formed in. ... 87
experiments in re-
roasting ore
from shaft of ... 107
experiments with . 31
in heap-roasting 28, 29
failure on ores of
Mexico 86
fuel needed in .... 43
in Bosque mill at
Parral 101,102
preferable to long
reverberatory ... 52
silicates formed in. 110
unsuitable for zinc
blende and ga-
lena ores Ill
used at Sombre-
rete, Zacatecas,
Mexico 254
used in roasting
* plumbiferous sil-
ver ores 41
used to roast zinc-
lead ore 104
Stir tanks 263
Storage tanks 298
for hyposulphite
solution 203
Straw, used for filter in leaching
tanks . , .182
INDEX
343
Sulphate of lime, formed in
roasting 8
Sulp hating roasting 94
time required for. 97
Sulphide minerals, classification
of, in relation to
roasting 9
ores 3
Sulphur in ores, effect of, on
quantity of fuel .... 42
lack of, overcome by
burning brimstone . . 86
removal of, by burning 207
by distillation 207
from silver precipi-
tate 203
used in preparing cal-
cium polysulphide . . 186
Sulphuric acid, anhydrous, for-
mation of .... 4
extraction with . 258
formed in sul-
phating roast-
ing 94
gas, formed in
roasting 4
Sulphurous acid converted into
anhydrous sul-
phuric acid 4
formed from ga-
lena in
roasting . . 6
in roasting . . 3
gas, formed in
heap-roasting 28,
29
chloride, formed in
roasting 3,6,7
Sump-tanks 298
Sustersic, F 8, 195, 197, 204
Sustersic's method of preparing
precipitate for refining 195
Tailing elevator-wheel 290
Tank, leaching, construction of 157
lixiviation 219
fineness of precipitate . 248
Tank lixiviation less advanta-
geous than trough lixi-
viation 249
quantity of solution re-
quired 248
time required for base-
metal leaching. 245
for silver leaching . . 248
Tarshish mine, Alpine county,
California 283, 285
Testing cyanide solution for gold
and silver 326
Tests for silver, in leaching. ... 178
Tools, best form of, for working
charges 52
Tower for refining cupric sul-
phate solutions 267
Trinidad, Mexico, lixiviation
at 155
Trough, distributing, for milk of
lime 188
lixiviation 181, 219
advantages of 249
arrangement and
operations of tanks 229
at Cusihuiriachic .... 240
fineness of precipi-
tate 248
precipitating vats. . . . 233
prevention of loss of
silver in 162
quantity of solution
required 248
settling-tank 225
silver dissolved by
base-metal solution
in 246
silver leaching 246
sluice-tanks and sluic-
ing 225
time required for base-
metal leaching .... 244
time required for silver
leaching 248
water required 245
Troughs, lixiviation 215
Two-story long furnace 60
344
INDEX
Two-story reverberatory furnace
compared with mod-
ified Howell 126
consumption of
wood in 125
cost of roasting in. 126
experiments with,
in roasting San
Francisco del Oro
ore 121
used in sulphating
roasting 95
single-hearth rever-
beratory furnace . . 46
Underfed stoker 272
United Zinc and Chemical Com-
pany, Argentine, Kansas .... 90
Veta Grande, Parral, Mexico,
leaching with cupric
chloride at 174
mine 101
Volatile chlorides, method of
avoiding expulsion of 20
Volatilization 20
loss of silver by . 117
Von Patera 155, 179
Waste liquor, recovery of silver
from 165
Water, consumption of, in trough
lixiviation 245
use of, to desilverize base-
metal chlorides 170
White-Howell furnace, at Parral,
Mexico 170
roasting zinc-
lead ore in . . 1 12
see also Howell-
White furnace
White lead, not to be used in
lixiviation troughs 215
Wilfley concentrators 290
Wood, amount required in roast-
ing zinc-lead ore 109
ashes, making lye from. 205
Wood, consumption of, in rever-
beratory furnace 125, 150
in roasting San Fran-
cisco del Oro ore ... 119
in self-roasting 141
used as fuel in roasting. 42
Woolly ore, cause of 13
Working doors, construction of,
in long reverberatory furnace 56
Yedras, Sinaloa, Mexico, behav-
ior of ore from ... 17
conclusions of experi-
ments with ores at 151
loss of silver in roast-
ing ore from 21
ores from 50
roasting calcareous
ores at 145
tests in loss of weight
in roasting ores
from 23
Ziervogel 94
process 94,281
Zinc blende, behavior of in
roasting 6,9
effect of salt on, in
roasting 16
steam on ores con-
taining 33
in ore from San
Francisco del Oro
mine 99
not favorable in
roasting in Stete-
feldt furnace ... 86
boxes 299,312
clean up of 300
chloride, formed in roasting 7
fumes, effect of, in
roasting 7
effect of, in ores 15
effect of, on time of lixivia-
tion 180
-lead ore, argentiferous, chlo-
ridizing of .... 99
INDEX
345
PAGE
Zinc-lead ore, chloridizing in
Howell furnace 82
from San Fran-
cisco del Oro
mine 28,31,61
PAGE
Zinc oxide, formed in roasting . 7
sulphate, formed in roast-
ing 7
sulphide, action of, in roast-
ing 6
YD 07559
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