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LEAD SMELTING

AND

REFINING

WITH SOME NOTES ON LEAD MINING

EDITED BY

WALTER RENTON INGALLS

— •

Published by the

McGraw-Hill Book Company

New Yoirk

Succe.s.sons to the 5ook Departments of the

McGraw Publishing Company Hill Publishing- Company

Publishers of Books for

Electrical World The Engineering and Mining Journal

Earing Record Amer;can ^.^

tJectric Railway Journal Coal Age

Metallurgical and Chemical Engineering power

LEAD SMELTING

AND

REFINING

WITH SOME NOTES ON LEAD MINING

EDITED BY

WALTER RENTON INGALLS

NEW YORK AND LONDON

THE ENGINEERING AND MINING JOURNAL
1906

/ & J

Is-

ALSO ENTERED AT
STATIONERS' HALL, LONDON, ENGLAND.

"* •% • . * '

"• V ':

PREFACE

THIS book is a reprint of various articles pertaining especially
to the smelting and refining of lead, together with a few articles
relating to the mining of lead ore, which have appeared in the
Engineering and Mining Journal, chiefly during the last three
years; in a few cases articles from earlier issues have been inserted,
in view of their special importance in rounding out certain of the
subjects treated. For the same reason, several articles from the
Transactions of the American Institute of Mining Engineers have
been incorporated, permission to republish them in this way
having been courteously granted by the Secretary of the Institute.
Certain of the other articles comprised in this book are abstracts
of papers originally presented before engineering societies, or
published in other technical periodicals, subsequently republished
in the Engineering and Mining Journal, as to which proper
acknowledgment has been made in all cases.

The articles comprised in this book relate to a variety of
subjects, which are of importance in the practical metallurgy of
lead, and especially in connection with the desulphurization of
galena, which is now accomplished by a new class of processes-
known as "Lime Roasting" processes. The successful introduc-
tion of these processes into the metallurgy of lead has been one
of the most important features in the history of the latter during
the last twenty-five years. Their development is so recent that
they are not elsewhere treated in technical literature, outside of
the pages of the periodicals and the transactions of engineering
societies. The theory and practice of these processes are not
yet by any means well understood, and a year or two hence we
shall doubtless possess much more knowledge concerning them
than we have now. Prompt information respecting such new
developments is, however, more desirable than delay with a view
to saying the last word on the subject, which never can be said
by any of us, even if we should wait to the end of the lifetime^

iii

337623

PREFACE

For this reason it has appeared useful to collect and republish
in convenient form the articles of this character which have
appeared during the last few years.

W. R. INGALLS.
AUGUST 1, 1906.

CONTENTS

PART I

NOTES ON LEAD MINING

PAGE

SOURCES OF LEAD PRODUCTION IN THE UNITED STATES (WALTER

RENTON INGALLS) 3

NOTES ON THE SOURCE OF THE SOUTHEAST MISSOURI LEAD (H. A.

WHEELER) 10

MINING IN SOUTHEASTERN MISSOURI (WALTER RENTON INGALLS) . . 16
LEAD MINING IN SOUTHEASTERN MISSOURI (R. D. O. JOHNSON) . .18
THE LEAD ORES OF SOUTHWESTERN MISSOURI (C. V. PETRAEUS AND

W. GEO. WARING 24

PART II
ROAST-REACTION SMELTING

SCOTCH HEARTHS AND REVERBERATORY FURNACES

LEAD SMELTING IN THE SCOTCH HEARTH (KENNETH W. M. MIDDLETON) 31

THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL. (O. PUFAHL) . . 38

LEAD SMELTING AT TARNOWITZ (EDITORIAL) 41

LEAD SMELTING IN REVERBERATORY FURNACES AT DESLOGE, Mo.

(WALTER RENTON INGALLS) __ 42

PART III

SINTERING AND BRIQUETTING

THE DESULPHURIZATION OF SLIMES BY HEAP ROASTING AT BROKEN

HILL (E. J. HORWOOD) 51

THE PREPARATION OF FINE MATERIAL FOR SMELTING (T. J. GREEN-
WAY) 59

THE BRIQUETTING OF MINERALS (ROBERT SCHORR) 63

A BRICKING PLANT FOR FLUE DUST AND FINE ORES (!AS. C. BENNETT) 66

PART IV

SMELTING IN THE BLAST FURNACE

MODERN SILVER-LEAD SMELTING (ARTHUR S. DWIGHT) 73

MECHANICAL FEEDING OF SILVER-LEAD BLAST FURNACES (ARTHUR S.

DWIGHT) 81

vi CONTENTS

PAGE

COST OF SMELTING AND REFINING (MALVERN W. ILES) 96

SMELTING ZINC RETORT RESIDUES (E. M. JOHNSON) 104

ZINC OXIDE IN SLAGS (W. MAYNARD HUTCHINGS) ...... 108

PART V

LlME-ROASTING OF GALENA

THE HUNTINGTON-HEBERLEIN PROCESS 113

LIME -ROASTING OF GALENA (EDITORIAL) 114

THE NEW METHODS OF DESULPHURIZING GALENA (W. BORCHERS) , 116

LlME-ROASTING OF GALENA (W. MAYNARD HUTCHINGS) .... 126

THEORETICAL ASPECTS OF LEAD-ORE ROASTING (C. GUILLEMAIN) . . 133
METALLURGICAL BEHAVIOR OF LEAD SULPHIDE AND CALCIUM SULPHATE

(F. O. DOELTZ) 139

THE HUNTINGTON-HEBERLEIN PROCESS (DONALD CLARK) .... 144
THE HUNTINGTON-HEBERLEIN PROCESS AT FRIEDRICHSHUTTE (A.

BIERNBAUM) 148

THE HUNTINGTON-HEBERLEIN PROCESS FROM THE HYGIENIC STAND-
POINT (A. BIERNBAUM) ...... 160

THE HUNTINGTON-HEBERLEIN PROCESS (THOMAS HUNTINGTON AND

FERDINAND HEBERLEIN) 167

MAKING SULPHURIC ACID AT BROKEN HILL (EDITORIAL) ... .174

THE CARMICHAEL-BRADFORD PROCESS (DONALD CLARK) 175

THB CARMICHAEL-BRADFORD PROCESS (WALTER RENTON INGALLS) . . 177

THE SAVELSBERG PROCESS (WALTER RENTON INGALLS) .... 186

LlME-ROASTING OF GALENA (WALTER RENTON INGALLS ) .... 193

PART VI

OTHER METHODS OF SMELTING

THE BORMETTES METHOD OF LEAD AND COPPER SMELTING (ALFBBDO

Lorn) 215

THE GERMOT PROCESS (WALTER RENTON INGALLS) ...... 224

PART VII

DUST AND FUME RECOVERY

FLUES, CHAMBERS AND BAG-HOUSES

DUST CHAMBER DESIGN (MAX J. WELCH) 229

CONCRETE IN METALLURGICAL CONSTRUCTION (HENRY W. EDWARDS) 234

CONCRETE FLUES (EDWIN H. MESSITER) 240

CONCRETE FLUES (FRANCIS T. HAVARD) 242

BAG-HOUSES FOR SAVING FUME (WALTER RENTON INGALLS) . . . 244

CONTENTS yii

PART VIII
BLOWERS AND BLOWING ENGINES

ROTARY BLOWERS vs. BLOWING ENGINES FOR LEAD SMELTING (Eoi- PAGE
TORIAL) 251

ROTARY BLOWERS vs. BLOWING ENGINES (J. PARKS CHANNING) . . 254

BLOWERS AND BLOWING ENGINES FOR LEAD AND COPPER SMELTING

(HIRAM W. HIXON) 256

BLOWING ENGINES AND ROTARY BLOWERS (S. E. BRETHERTON) . . 258

PART IX
LEAD REFINING

THE REFINING OF LEAD BULLION (F. L. PIDDINGTON) 263

THE ELECTROLYTIC REFINING OF BASE LEAD BULLION (Trrus ULKE) 270
ELECTROLYTIC LEAD REFINING (ANSON G. BETTS) 274

PART X

SMELTING WORKS AND REFINERIES

THE NEW SMELTER AT EL PASO, TEXAS (EDITORIAL) 365

NEW PLANT OF THE AMERICAN SMELTING AND REFINING COMPANY AT

MURRAY, UTAH (WALTER RENTON INGALLS) . . . . . . 287

THE MURRAY SMELTER, UTAH (O. PUFAHL) 291

THE PUEBLO LEAD SMELTERS (O. PUFAHL) 294

THE PERTH AMBOY PLANT OF THE AMERICAN SMELTING AND REFINING

COMPANY (O. PUFAHL) . . 296

THE NATIONAL PLANT OF THE AMERICAN SMELTING AND REFINING

COMPANY (O. PUFAHL) 299

THE EAST HELENA PLANT OF THE AMERICAN SMELTING AND REFINING

COMPANY (O. PUFAHL) 302

THE GLOBE PLANT OF THE AMERICAN SMELTING AND REFINING COM-
PANY (O. PUFAHL) 304

LEAD SMELTING IN SPAIN (Hj ALMAR ERIKSSON) 306

LEAD SMELTING AT MONTEPONI, SARDINIA (ERMINIO FERRARIS) . .311

PART I
NOTES ON LEAD MINING

SOURCES OF LEAD PRODUCTION IN THE UNITED

STATES

BY WALTER RENTON INGALLS

(November 28, 1903)

Statistics of lead production are of value in two directions:
(1) in showing the relative proportion of the kinds of lead pro-
duced; and (2) in showing the sources from which produced. Lead
is marketed in three principal forms: (a) desilverized; (6) soft; (c)
antimonial, or hard. The terms to distinguish between classes
"a" and "b" are inexact, because, of course, desilverized lead is
soft lead. Desilverized lead itself is classified as " corroding/'
which is the highest grade, and ordinary "desilverized." Soft
lead, referring to the Missouri product, may be either "ordinary"
or "chemical hard." The latter is such lead as contains a small
percentage of copper and antimony as impurities, which, without
making it really hard, increase its resistance against the action
of acids, and therefore render it especially suitable for the pro-
duction of sheet to be used in sulphuric-acid chamber construction
and like purposes. The production of chemical hard lead is a
fortuitous matter, depending on the presence of the valuable
impurities in the virgin ores. If present, these impurities go
into the lead, and cannot be completely removed by the simple
process of refining which is practised. Nobody knows just what
proportions of copper and antimony are required to impart the
desired property, and consequently no specifications are made.
Some chemical engineers call for a particular brand, but this is
really only a whim, since the same brand will not be uniformly
the same; practically one brand is as good as another. Corroding
lead is the very pure metal, which is suitable for white lead
manufacture. It may be made either from desilverized or from
the ordinary Missouri product; or the latter, if especially pure,
may be classed as corroding without further refining. Antimonial
lead is really an alloy of lead with about 15 to 30 per cent, anti-
mony, which is produced as a by-product by the desilverizers of

4 LEAD SMELTING AND REFINING

base bullion. The antimony content is variable, it being possible
for the smelter to run the percentage up to 60. Formerly it was
the general custom to make antimonial lead with a content of
10 to 12 per cent. Sb; later, with 18 to 20 per cent.; while now
25 to 30 per cent. Sb is best suited to the market.

The relative values of the various grades of lead fluctuate
considerably, according to the market place, and the demand and
supply. The schedules of the American Smelting and Refining
Company make a regular differential of lOc. per 100 Ib. between
corroding lead and desilverized lead in all markets. In the
St. Louis market, desilverized lead used to command a premium
of 5c. to lOc. per 100 Ib. over ordinary Missouri; but now they
sell on approximately equal terms. Chemical hard lead sells
sometimes at a higher price, sometimes at a lower price, than
ordinary Missouri lead, according to the demand and supply.
There is no regular differential. This is also the case with anti-
monial lead.1

The total production of lead from ores mined in the United
States in 1901 was 279,922 short tons, of which 211,368 tons
were desilverized, 57,898 soft (meaning lead from Missouri and
adjacent States) and 10,656 antimonial. These are the statistics
of " The Mineral Industry." The United States Geological Survey
reported substantially the same quantities. In 1902 the pro-
duction was 199,615 tons of desilverized, 70,424 tons of soft,
and 10,485 tons of antimonial, a total of 280,524 tons. There is
an annual production of 4000 to 5000 tons of white lead direct
from ore at Joplin, Mo., which increases the total lead production
of the United States by, say, 3500 tons per annum. The produc-
tion of lead reported as "soft" does not represent the full output
of Missouri and adjacent States, because a good deal of their
ore, itself non-argentiferous, except to the extent of about 1 oz.
per ton in certain districts, is smelted with silver-bearing ores,
going thus into an argentiferous lead; while in one case, at least,
the almost non-argentiferous lead, obtained by smelting the ore
unmixed, is desilverized for the sake of the extra refining.

Lead-bearing ores are of widespread occurrence in the United
States. Throughout the Rocky Mountains there are numerous
districts in which the ore carries more or less lead in connection

1 During 1905, antimonial lead commanded a premium of about Ic.
per Ib. above desilverized, owing to the high price for antimony.

NOTES ON LEAD MINING 5

with gold and silver. For this reason, the lead mining industry is
not commonly thought of as having such a concentrated char-
acter as copper mining and zinc mining. It is the fact, however,
that upward of 70 per cent, of the lead produced in the United
States is derived from five districts, and in the three leading
districts from a comparatively small number of mines. The
statistics of these for 1901 to 1904 are as follows: 1

DISTRICT

^ ^ ^

Tt~~

SENT.—

1903

1904

1901

1902

1903

1904

1901

1902

Coeur d'Alene. . .
Southeast Mo. . .
Leadville,Colo. .
Park City, Utah
Joplin , Mo .-Kan .

Total

68,953
46,657
28,180
28,310
24,500

74,739
56,550
19,725
36,300
22,130

89,880
59,660
18,177
36,534
20,000

98,240
59,104
23,590
30,192
23,600

24.3
16.4
10.0
10.0
8.6

26.3
19.9
6.9
12.8
7.8

32.5
21.2
6.6
13.2

7.2

32.5
19.6

7.8
10.0

7.8

a
b
c
d

196,600

209,444

224,251

234,726

69.3

73.7

81.0

77.7

a. The production in 1901 and 1902 is computed from direct returns from
the mines, with an allowance of 6 per cent, for loss of lead in smelting. The
production in 1903 and 1904 is estimated at 95 per cent, of the total lead
product of Idaho.

6. This figure includes only the output of the mines at Bonne Terre, Flat
River, Doe Run, Mine la Motte and Fredericktown. It is computed from the
report of the State Lead and Zinc Mine Inspector as to ore produced, the ore
(concentrates) of the mines at Bonne Terre, Flat River and Doe Run being
reckoned as yielding 60 per cent. lead.

c. Report of State Commissioner of Mines.

d. Report of Director of the Mint on "Production of Gold and Silver in
the United States," with allowance of 6 per cent, for loss of lead in smelting.

e. From statistics reported by "The Mineral Industry," reckoning the ore
(concentrates) as yielding 70 per cent. lead.

Outside of these five districts, the most of the lead produced
in the United States is derived from other camps in Idaho, Colo-
rado, Missouri and Utah, the combined output of all other States
being insignificant. It is interesting to examine the conditions
under which lead is produced in the five principal districts.

Leadville, Colo. — The mines of Leadville, which once were the
largest lead producers of the United States, became comparatively
unimportant after the exhaustion of the deposits of carbonate
ore, but have attained a new importance since the successful

1 The figures for 1903 and 1904 have been added in the revision of this
article for this book. The production of lead in the United States in 1903
was 276,694 tons; in 1904, it was 302,204 tons.

6 LEAD SMELTING AND REFINING

introduction of means for separating the mixed sulphide ore,
which occurs there in very large bodies. The lead production of
Leadville in 1897 was 11,850 tons; 17,973 tons in 1898; 24,299
tons in 1899; 31,300 tons in 1900; 28,180 tons in 1901, and 19,725
tons in 1902. The Leadville mixed sulphide ore assays about
8 per cent. Pb, 25 per cent. Zn and 10 oz. silver per ton. It is sep-
arated into a zinc product assaying about 38 per cent. Zn and 6 per
cent. Pb, and a galena product assaying about 45 per cent. Pb,
10 or 12 per cent. Zn, and 10 oz. silver per ton.

Coeur d'Alene. — The mines of this district are opened on
fissure veins of great extent. The ore is of low grade and requires
concentration. As mined, it contains about 10 per cent, lead
and a variable proportion of silver. It is marketed as mineral,
averaging about 50 per cent. Pb and 30 oz. silver per ton. The
production of lead ore in this district is carried on under the
disadvantages of remoteness from the principal markets for pig
lead, high-priced labor, and comparatively expensive supplies.
It enjoys the advantages of large orebodies of comparatively
high grade in lead, and an important silver content, and in many
cases cheap water power, and the ability to work the mines
through adit levels. The cost of mining and milling a ton of
crude ore is $2.50 to $3.50. The mills are situated, generally,
at some distance from the mines, the ore being transported by
railway at a cost of 8 to 20c. per ton. The dressing is done in
large mills at a cost of 40 to 50c. per ton. About 75 per cent, of
the lead of the ore is recovered. The concentrates are sold at
90 per cent, of their lead contents and 95 per cent, of their silver
contents, less a smelting charge of $8 per ton, and a freight rate
of $8 per ton on ore of less than $50 value per ton, $10 on ore
worth $50 to $65, and $12 on ore worth more than $65; the ore
values being computed f. o. b. mines. The settling price of lead
is the arbitrary one made by the American Smelting and Refining
Company. With lead (in ore) at 3.5c. and silver at 50c., the
value, f. o. b. mines, of a ton of ore containing 50 per cent. Pb
and 30 oz. silver is approximately as follows:

1000 X 0.90 = 900 Ib. lead, at 3.5c $31.50

30 X 0.95 - 28.5 oz. silver, at 50c 14.25

Gross value, f. o. b. mines $45.75

Less freight, $10, and smelting charge, $8 18.00

Net value, f. o. b. mines $27.75

NOTES ON LEAD MINING

Assuming an average of 6 tons of crude ore to 1 ton of con-
centrate, the value per ton of crude ore would be about $4.62J,
and the net profit per ton about $1.62$, which figures are increased
23.75c. for each 5c. rise in the value of silver above 50c. per
ounce.

The production of the Cceur d'Alene since 1895, as reported
by the mines, has been as follows:

YEAR

LEAD, TONS

SILVER, oz.

RATIO1

1896

37250

2 500000

67 1

1897

57777

3 579 424

61 9

1898

56 339

3 399 524

60 3

1899

50006

2 736 872

54 7

1900

81 535

4 755 877

583

1901

68953

3 349 533

485

1902 . ...

74 739

4 489 549

600

1903

2 100 355

5*751 613

573

1904

2 108 954

6 247 795

574

The number of producers in the Coeur d'Alene district is
comparatively small, and many of them have recently consoli-
dated, under the name of the Federal Mining and Smelting
Company. Outside of that concern are the Bunker Hill &
Sullivan, the Morning and the Hercules mines, control of which
has lately been secured by the American Smelting and Refining
Company.

Southeastern Missouri. — The most of the lead produced in
this region comes from what is called the disseminated district,
comprising the mines of Bonne Terre, Flat River, Doe Run,
Mine la Motte and Fredericktown, of which those of Bonne Terre
and Flat River are the most important. The ore of this region
is a magnesian limestone impregnated with galena. The deposits
lie nearly flat and are very large. They yield about 5 per cent, of
mineral, which assays about 65 per cent. lead. The low grade of
the ore is the only disadvantage which this district has, but this is
so much more than offset by the numerous advantages, that mining
is conducted very profitably, and it is an open question whether
lead can be produced more cheaply here or in the Coeur d'Alene.
The mines of southeastern Missouri are only 60 to 100 miles

1 Ounces of silver to the ton of lead.

2 These figures are doubtful; they are probably too high. (See table on
p. 5).

8 LEAD SMELTING AND REFINING

distant from St. Louis, and are in close proximity to the coal-
fields of southern Illinois, which afford cheap fuel. The ore lies
at depths of only 100 to 500 ft. below the surface. The ground
stands admirably, without any timbering. Labor and supplies
are comparatively cheap. Mining and milling can be done for
$1.05 to $1.25 per ton of crude ore, when conducted on the large
scale that is common in this district. Most of the mining com-
panies are equipped to smelt their own ore, the smelters being
either at the mines or near St. Louis. The freight rate on con-
centrates to St. Louis is $1.40 per ton; on pig lead it is $2.10 per
ton. The total cost of producing pig lead, delivered at St. Louis,
is about 2.25c. per pound, not allowing for interest on the invest-
ment, amortization, etc.

The production of the mines in the disseminated district in
1901 was equivalent to 46,657 tons of pig lead; in 1902 it was
56,550 tons. The milling capacity of the district is about 6000
tons per day, which corresponds to a capacity for the production
of about 57,000 tons of pig lead per annum. The St. Joseph
Lead Company is building a new 1000-ton mill, and the St. Louis
Smelting and Refining Company (National Lead Company) is
further increasing its output, which improvements will increase
the daily milling capacity by about 1400 tons, and will enable
the district to put out upward of 66,000 tons of pig lead. In
this district, as in the Cceur d'Alene, the industry is closely
concentrated, there being only nine producers, all told.

Park City, Utah. — Nearly all the lead produced by this
camp comes from the Silver King, Daly West, Ontario, Quincy,
Anchor and Daly mines, which have large veins of low-grade ore
carrying argentiferous galena and blende, a galena product being
obtained by dressing, and zinkiferous tailings, which are accu-
mulated for further treatment as zinc ore, when market conditions
justify.1

Joplin District. — The lead production of southwestern Mis-
souri and southeastern Kansas, in what is known as the Joplin
district, is derived entirely as a by-product in dressing the zinc
ore of that district. It is obtained as a product assaying about
77 per cent. Pb, and is the highest grade of lead ore produced,
in large quantity, anywhere in the United States. It is smelted
partly for the production of pig lead, and partly for the direct

1 The production of zinc ore in this district has now been commenced.

NOTES ON LEAD MINING 9

manufacture of white lead. The lead ore production of the
district was 31,294 tons in 1895, 26,927 tons in 1896, 29,578 tons
in 1897, 26,457 tons in 1898, 24,100 tons in 1899, 28,500 tons in
1900, 35,000 tons in 1901, and 31,615 tons in 1902. The pro-
duction of lead ore in this district varies more or less, according
to the production of zinc ore, and is not likely to increase mate-
rially over the figure attained in 1901.

NOTES ON THE SOURCE OF THE SOUTHEAST
MISSOURI LEAD

BY H. A. WHEELER

(March 31, 1904)

The source of the lead that is being mined in large quantities
in southeastern Missouri has been a mooted question. Nor is
the origin of the lead a purely theoretical question, as it has an
important bearing on the possible extension of the orebodies
into the underlying sandstone.

The disseminated lead ores of Missouri occur in a shaly,
magnesian limestone of Cambrian age in St. Francois, Madison
and Washington counties, from 60 to 130 miles south of St. Louis.
The limestone is known as the Bonne Terre, or lower half of
"the third magnesian limestone" of the Missouri Geological
Survey, and rests on a sandstone, known as " the third sandstone,"
that is the base of the sedimentary formations in the area. Under
this sandstone occur the crystalline porphyries and granites of
Algonkian and Archean age, which outcrop as knobs and islands
of limited extent amid the unaltered Cambrian and Lower Silurian
sediments.

The lead occurs as irregular granules of galena scattered
through the limestone in essentially horizontal bodies that vary
from 5 to 100 ft. in thickness, from 25 to 500 ft. in width, and
have exceeded 9000 ft. in length. There is no vein structure, no
crushing or brecciation of the inclosing rock, yet these orebodies
have well defined axes or courses, and remarkable reliability and
persistency. It is true that the limestone is usually darker,
more porous, and more apt to have thin seams of very dark
(organic) shales where it is ore-bearing than in the surrounding
barren ground. The orebodies, however, fade out gradually,
with no sharp line between the pay-rock and the non-paying,
and the lead is rarely, if ever, entirely absent in any extent of
the limestone of the region. While the main course of the ore-
bodies seems to be intimately connected with the axes of the

NOTES ON LEAD MINING 11

gentle anticlinal folds, numerous cross-runs of ore that are asso-
ciated with slight faults are almost as important as the main
shoots, and have been followed for 5000 ft. in length. These
cross-runs are sometimes richer than the main runs, at least
near the intersections, but they are narrower, and partake more
of the type of vertical shoots, as distinguished from the horizontal
sheet-form.

Most of the orebodies occur at, or close to, the base of the
limestone, and frequently in the transition rock between the
underlying sandstone and the limestone, though some notable
and important bodies have been found from 100 to 200 ft. above
the sandstone. This makes the working depth from the surface
vary from 150 to 250 ft., for the upper orebodies, to 300 to 500 ft.
deep to the main or basal orebodies, according as erosion has
removed the ore-bearing limestone. The thickness of the latter
ranges from 400 to 500 ft.

Associated with the galena are less amounts of pyrite, which
especially fringes the orebodies, and very small quantities of
chalcopyrite, zinc blende, and siegenite (the double sulphide of
nickel and cobalt). Calcite also occurs, especially where recent
leaching has opened vugs, caves, or channels in the limestone,
when secondary enrichment frequently incrusts these openings
with crystals of calcite and galena. No barite ever occurs with
the disseminated ore, though it is the principal gangue mineral
in the upper or Potosi member of the third magnesian limestone,
and is never absent in the small ore occurrences in the still higher
second magnesian limestone.

While the average tenor of the ore is low, the yield being from
3 to 4 per cent, in pig lead, they are so persistent and easy to
mine that the district today is producing about 70,000 tons of
pig lead annually, and at a very satisfactory profit. As the
output was about 2500 tons lead in 1873, approximately 8500
tons in 1883, and about 20,000 tons in 1893, it shows that this
district is young, for the principal growth has been within the
last five years.

Of the numerous but much smaller occurrences of lead else-
where in Missouri and the Mississippi valley, none resembles this
district in character, a fact which is unique. For while the
Mechernich lead deposits, in Germany, are disseminated, and of
even lower grade than in Missouri, they occur in a sandstone,

12 LEAD SMELTING AND REFINING

and (like all the lead deposits outside of the Mississippi valley)
they are argentiferous, at least to an extent sufficient to make
the extraction of the silver profitable; and on the non-argentiferous
character of the disseminated deposits hangs my story.

Of the numerous hypotheses advanced to account for the
origin of these deposits, there are only two that seem worthy of
consideration: (a) the lateral secretion theory, and (b) deposition
from solutions of deep-seated origin. Other theories evolved in
the pioneer period of economic geology are interesting, chiefly by
reason of the difficulties under which the early strugglers after
geological knowledge blazed the pathway for modern research
and observation.

The lateral secretion theory, as now modernized into the
secondary enrichment hypothesis, has much merit when applied
to the southeastern and central Missouri lead deposits. For the
limestones throughout Missouri — and they are the outcropping
formation over more than half of the State — are rarely, if ever,
devoid of at least slight amounts of lead and zinc, although they
range in age from the Carboniferous down to the Cambrian.

The sub-Carboniferous formation is almost entirely made up
of limestones, which aggregate 1200 to 1500 ft. in thickness.
They frequently contain enough lead (and less often zinc) to
arouse the hopes of the farmer, and more or less prospecting has
been carried on from Hannibal to St. Louis, or 125 miles along
the Mississippi front, and west to the central part of the State,
but with most discouraging results.

In the rock quarries of St. Louis, immediately under the
lower coal measures, fine specimens of millerite of world-wide
reputation occur as filiform linings of vugs in this sub-Carbonif-
erous limestone. These vugs occur in a solid, unaltered rock
which gives no clue to the existence of the vug or cavity until it
is accidentally broken. The vugs are lined with crystals of pink
dolomite, calcite and millerite, with occasionally barite, selenite,
galena and blende. They occur in a well-defined horizon about
5 ft. thick, and the vugs in the limestone above and below this
millerite bed contain only calcite, or less frequently dolomite.
Yet this sub-Carboniferous formation in southwestern Missouri,
about Joplin, carries the innumerable pockets and sheets of lead
and zinc that have made that district the most important zinc
producer in the world. While faulting and limited folding occur

NOTES ON LEAD MINING 13

in eastern and central Missouri to fully as great an extent as in
St. Frangois county or the Joplin district, thus far no mineral
concentration into workable orebodies has been found in this
formation, except in the Joplin area.

The next important series of limestones that make up most
of the central portion of Missouri are of Silurian age, and in them
lead and zinc are liberally scattered over large areas. In the
residual surface clays left by dissolution of the limestone, the
farmers frequently make low wages by gophering after the liber-
ated lead, and the aggregate of these numerous though insignificant
gopher-holes makes quite a respectable total. But they are only
worked when there is nothing else to do on the farm, as with rare
exceptions they do not yield living wages, and the financial
results of mining the rock are even less satisfactory. Yet a few
small orebodies have been found that were undoubtedly formed
by local leaching and re-precipitation of this diffused lead and
zinc. Such orebodies occur in openings or caves, with well
crystallized forms of galena and blende, and invariably associated
with crystallized "tiff" or barite. I am not aware of any of
these pockets or secondary enrichments having produced as much
as 2000 tons of lead or zinc, and very few have produced as
much as 500 tons, although one of these pockets was recently
exploited with heroic quantities of printer's ink as the largest
lead mine in the world. Yet there are large areas in which it is
almost impossible to put down a drill-hole without finding
" shines" or trifling amounts of lead or zinc. That these central
Missouri lead deposits are due to lateral secretion there seems,
little doubt, and it is possible that larger pockets may yet be
found where more favorable conditions occur.

When the lateral secretion theory is applied to the dissemi-
nated deposits of southeastern Missouri, we are confronted by
enormous bodies of ore, absence of barite, non-crystallized condi-
tion of the galena except in local, small, evidently secondary
deposits, and well-defined courses for the main and cross-runs of
ore. The Bonne Terre orebody, which has been worked longest
and most energetically, has attained a length of nearly 9000 ft.,
with a production of about 350,000 tons or $30,000,000 of lead,
and is far from being exhausted. Orebodies recently opened are
quite as promising. The country rock is not as broken nor as
open as in central Missouri, and is therefore much less favorable

14 LEAD SMELTING AND REFINING

for the lateral circulation of mineral waters, yet the orebodies
vastly exceed those of the central region.

Further, the Bonne Terre formation is heavily intercalated
with thick sheets of shale that would hinder overlying waters
from reaching the base of the ore-horizon, where most of the ore
occurs, so that the leachable area would be confined to a very
limited vertical range, or to but little greater thickness than the
100 ft. or so in which most of the orebodies occur. While I have
always felt that such large bodies, showing relatively rapid
precipitation of the lead, could not be satisfactorily explained
except as having a deep-seated origin, the fact that the dissemi-
nated ore is practically non-argentiferous, or at least carries only
one to three ounces per ton, has been a formidable obstacle.
For the lead in the small fissure-veins that occasionally occur in
the adjacent granite has always been reported as argentiferous.
Thus the Einstein silver mine, near Fredericktown, worked a
fissure-vein from 1 to 6 ft. wide in the granite. It had a typical
complex vein-filling and structure, and carried galena that assayed
from 40 to 200 oz. per ton. While the quantity of ore obtained
did not justify the expensive plant erected to operate it, the
galena was rich in silver, whereas in the disseminated ores at the
Mine la Motte mine, ten miles distant, only the customary 1.5 oz.
per ton occurs. Occasionally fine-grained specimens of galena
that I have found in the disseminated belt would unquestionably
be rated as argentiferous by a Western miner, but the .assay
showed that the structure in this case was due to other causes,
as only about two ounces were found. An apparent exception
was reported at the Peach Orchard diggings, in Washington
county, in the higher or Potosi member of the third magnesian
limestone, where Arthur Thacher found sulphide and carbonate
ore carrying 8 to 10 oz. of silver per ton; and a short-lived hamlet,
known as Silver City, sprang up to work them. I found, however,
that these deposits are associated with little vertical fissure-veins
or seams that unquestionably come up from the underlying
porphyry.

Recently I examined the Jackson Revel mine, which has been
considered a silver mine for the last fifty years. It lies about
seven miles south of Fredericktown, and is a fissure-vein in
Algonkian felsite, where it protrudes, as a low hill, through the
disseminated limestone formation. A shaft has just been sunk

NOTES ON LEAD MINING 15

about 150 ft. at less than 1000 ft. from the feather edge of the
limestone. The vein is narrow, only one to twelve inches wide,
with slicken-sided walls, runs about N. 20 deg. E., and dips
80 to 86 deg. eastward. White quartz forms the principal part
of the filling; the vein contains more or less galena and zinc blende.
Assays of the clean galena made by Prof. W. B. Potter show only
2.5 oz. silver per ton, or no more than is frequently found in the
disseminated lead ores. As the lead in this fissure vein may be
regarded as of undoubted deep origin, and it is practically non-
argentiferous, this would seem to remove the last objection to
the theory of the deep-seated source of the lead in the disseminated
deposits of southeast Missouri.

MINING IN SOUTHEASTERN MISSOURI

BY WALTER RENTON INGALLS

(February 18, 1904)

The St. Joseph Lead Company, in the operation of its mines
at Bonne Terre, does not permit the cages employed for hoisting
purposes to be used for access to the mine. Men going to and
from their work must climb the ladders. This rule does not
obtain in the other mines of the district. The St. Joseph Lead
Company employs electric haulage for the transport of ore under-
ground at Bonne Terre. In the other mines of the district, mules
are generally used. The flow of water in the mines of the district
is extremely variable; some have very little; others have a good
deal. The Central mine is one of the wettest in the entire district,
making about 2000 gal. of water per minute. Coal in south-
eastern Missouri costs $2 to $2.25 per ton delivered at the mines,
and the cost of raising 2000 gal. of water per minute from a depth
of something like 350 ft. is a very considerable item in the cost
of mining and milling, which, in the aggregate, is expected to
come to not much over $1.25 per ton.

The ore shoots in the district are unusually large. Their
precise trend has not been identified. Some consider the pre-
dominance of trend to be northeast; others, northwest. They
go both ways, and appear to make the greatest depositions of
ore at their intersections. However, the network of shoots, if
that be the actual occurrence, is laid out on a very grand scale.
Vertically there is also a difference. Some shafts penetrate only
one stratum of ore; others, two or three. The orebody may be
only a few feet in thickness; it may be 100 ft. or more. The
occurrence of several overlying orebodies obviously indicates
the mineralization of different strata of limestone, while in the
very thick orebodies the whole zone has apparently been miner-
alized.

The grade of the ore is extremely variable. It may be only
1 or 2 per cent, mineral, or it may be 15 per cent, or more. How-

NOTES ON LEAD MINING 17

ever, the average yield for the district, in large mines which
mill 500 to 1200 tons of ore per day, is probably about 5 per
cent, of mineral, assaying 65 per cent. Pb, which would correspond
to a yield of 3.25 per cent, metallic lead in the form of concentrate.
The actual recovery in the dressing works is probably about
75 per cent., which would indicate a tenure of about 4.33 per
cent, lead in the crude ore.

LEAD MINING IN SOUTHEASTERN MISSOURI
BY R. D. O. JOHNSON

(September 16, 1905)

The lead deposits of southeastern Missouri carry galena dis-
seminated in certain strata of magnesian limestone. Their greater
dimensions are generally horizontal, but with outlines extremely
irregular. The large orebodies consist usually of a series of
smaller bodies disposed parallel to one another. These smaller
members may coalesce, but are generally separated from one
another by a varying thickness of lean ore or barren rock. The
vertical and lateral dimensions of an orebody may be determined
with a fair degree of accuracy by diamond drilling, and a map
may be constructed from the information so obtained. Such a
map (on which are plotted the surface contours) makes it possible
to determine closely the proper location of the shaft, or shafts,
considering also the surface and underground drainage and
tramming.

The first shafts in the district were sunk at Bonne Terre,
where the deposits lie comparatively near the surface. The .early
practice at this point was to sink a number of small one-com-
partment shafts. As the deposits were followed deeper, this
gave way to the practice of putting down two-compartment
shafts equipped much more completely than were the shallower
shafts.

At Flat River (where the deposits lie at much greater depths,
some being over 500 ft.) the shafts are 7 x 14 ft., 6J x 18 ft.,
and 7 x 20 ft. These larger dimensions give room not only
for two cage- ways and a ladder-way, but also for a roomy pipe-
compartment. The large quantities of water to be pumped in
this part of the district make the care of the pipes in the shafts
a matter of first importance. At Bonne Terre only such a quan-
tity of water was encountered as could be handled by bailing or
be taken out with the rock; there the only pipe necessary was a
small air-pipe down one comer of the shaft. When the quantity

NOTES ON LEAD MINING 19

of water encountered is so great that the continued working of
the mine depends upon its uninterrupted removal, the care of
the pipes becomes a matter of great importance. A shaft which
yields from 4000 to 5000 gal. of water per minute is equipped
with two 12-in. column pipes and two 4-in. steam pipes covered
and sheathed. Moreover, the pipe compartment will probably
contain an 8-in. air-pipe, besides speaking-tubes, pipes for carrying
electric wires, and pipes for conducting water from upper levels
to the sump. To care for these properly there are required a
separate compartment and plenty of room.

Shafts are sunk by using temporary head frames and iron
buckets of from 8 to 14 cu. ft. capacity. Where the influx of
water was small, 104 ft. have been sunk in 30 days, with three
8-hour shifts, two drills, and two men to each drill; 2j-in. drills
are used almost exclusively; 3J-in. drills have been used in sinking,
but without apparent increase in speed.

The influence of the quantity of water encountered upon the
speed of sinking (and the consequent cost per foot) is so great
that figures are of little value. Conditions are not at all uniform.

At some point (usually before 200 ft. is reached) a horizontal
opening will be encountered. This opening invariably yields
water, the amount following closely the surface precipitation.
It is the practice to establish at this point a pumping station.
The shaft is " ringed" and the water is directed into a sump on
the side, from which it is pumped out. This sump receives also
the discharge of the sinking pumps.

The shafts sunk in solid limestone require no timbering other
than that necessary to support the guides, pipes, and ladder
platforms. These timbers are 8 x 8 in. and 6 x 8 in., spaced
7 or 8 ft. apart.

Shafts are sunk to a depth of 10 ft. below the point determined
upon as the lower cage landing. From the end at the bottom a
narrow drift is driven horizontally to a distance of 15 ft.; at that
point it is widened out to 10 ft. and driven 20 ft. further. The
whole area (10 x 20 ft.) is then rasied to a point 28 or 30 ft.
above the bottom of the drift from the shaft. The lower part of
this chamber constitutes the sump. Starting from this chamber
(on one side and at a point 10 ft. above the cage landing, or
20 ft. above the bottom of the sump), the "pump-house" is cut
out. This pump-house is cut 40 ft. long and is as wide as the

20 LEAD SMELTING AND REFINING

sump is long, namely, 20 ft. A narrow drift is driven to connect
the top of the pump-house with the shaft. Through this drift
the various pipes enter the pump-house from the shaft.

The pumps are thus placed at an elevation of 10 ft. above the
bottom of the mine. Flooding of mines, due to failure of pumps
or to striking underground bodies of water, taught the necessity
of placing the pumps at such an elevation that they would be the
last to be covered, thus giving time for getting or keeping them
in operation. The pumps are placed on the solid rock, the air-
pumps and condensers at a lower level on timbers over the sump.

With this arrangement, the bottom of the shaft serves as an
antechamber for the sump, in which is collected the washing
from the mine and the dripping from the shaft. The sump
proper rarely needs cleaning.

The pumps are generally of high-grade, compound- and triple-
expansion, pot-valved, outside-packed plunger pattern. Plants
with electrical power distribution have recently installed direct-
connected compound centrifugal pumps with entire success.

Pumps of the Cornish pattern have never been used much in
this region. One such pump has been installed, but the example
has not been followed even by the company putting it in.

The irregular disposition of the ore renders any systematic
plan of drifting or mining (as in coal or vein mining) impossible.
On each side of the shaft and in a direction at right angles to its
greater horizontal dimension, drifts 12 to 14 ft. in width are
driven to a distance of 60 or 70 ft. In these broad drifts are
located the tracks and also the " crossovers" for running the cars
on and off the cage.

When a deposit is first opened up, it is usually worked on
two, and sometimes three, levels. These eventually cut into one
another, when the ore is hoisted from the lower level alone.

The determination of the depth of the lower level is a matter
of compromise. Much good ore may be known to exist below;
when it comes to mining, it will have to be taken out at greater
expense; but the level is aimed to cut through the lower portions
of the main body. It is generally safe to predict that the ore
lying below the upper levels will eventually be mined from a
lower level without the expense of local underground hoisting
and pumping.

The ore has simply to be followed; no one can say in advance

NOTES ON LEAD MINING 21

how it is going to turn out. The irregularity of the deposits
renders any general plan of mining of little or no value. Some
managers endeavor to outline the deposits by working on the
outskirts, leaving the interior as "ore reserves." Such reserves
have been found to be no reserves at all, though the quality of
the rock may be fairly well determined by underground diamond
drilling. Many of the deposits are too narrow to permit the
employment of any system of outlining and at the same time
keeping up the ore supply.

The individual bodies constituting the general orebody are
rarely, if ever, completely separated by barren rock; some
" stringers" or "leaders" of ore connect them. The careful
superintendent keeps a record on the monthly mine map of all
such occurrences, or otherwise, or of blank walls of barren rock
that mark the edge of the deposit. This precaution finds abun-
dant reward when the drills commence to "cut poor," and when
a search for ore is necessary.

The method of mining is to rise to the top of the ore and to
carry forward a 6-ft. breast. If the ore is thick enough, this is
followed by the underhand stope. Drill holes in the breast are
usually 7 or 8 ft. in depth; stope holes, 10 to 14 feet.

Both the roof and the floor are drilled with 8- or 10-ft. holes
placed 8 or 10 ft. apart. These serve to prospect the rock in the
immediate neighborhood; in the roof they serve the further very
important purpose of draining out water that might otherwise
accumulate between the strata and that might force them to fall.
The condition or safety of the roof is determined by striking with
a hammer. If the sound is hollow or "drummy," the roof is
unsafe. If water is allowed to accumulate between the loose
strata, obviously it is not possible to determine the condition of
the roof.

It is the duty of two men on each shift to keep the mine in a
safe condition by taking down all loose and dangerous masses of
rock. These men are known as " miners." It sometimes happens
that a considerable area of the roof gets into such a dangerous
condition that it is either too risky or too expensive to put in
order, in which case the space underneath is fenced off. As a
general thing, the mines are safe and are kept so. There are but
few accidents of a serious nature due to falling rock.

The roof is supported entirely by pillars; no timbering what-

22 LEAD SMELTING AND REFINING

ever is used. The pillars are parts of the orebody or rock that is
left. They are of all varieties of size and shape. They are
usually circular in cross-section, 10 to 15 ft. in diameter and
spaced 20 to 35 ft. apart, depending upon the character of the
roof. Pillars generally flare at the top to give as much support
to the roof as possible. The hight of the pillars corresponds, of
course, to the thickness of the orebody.

All drilling is done by 2f-in. percussion drills. In the early
days, when diamonds were worth $6 per carat, underground
diamond drills were used. Diamond drills are used now occa-
sionally for putting in long horizontal holes for shooting down
"drummy" roof. Air pressure varies from 60 to 80 Ib. Pres-
sures of 100 Ib. and more have been used, but the repairs on the
drills became so great that the advantages of the higher pressure
were neutralized.

Each drill is operated by two men, designated as "drillers,"
or "front hand" and "back hand." The average amount of
drilling per shift of 10 hours is in the neighborhood of 40 ft.,
though at one mine an average of 55 ft. was maintained.

In some of the mines the "drillers" and "back hands" do the
loading and firing; in others that is done by "firers," who do the
blasting between shifts. When the drillers do the firing, there is
employed a "powder monkey," who makes up the "niphters,"
or sticks of powder, in which are inserted and fastened the caps
and fuse; 35 per cent, powder is used for general mining.

Battery firing is employed only in shaft sinking. In the
mining work this is found to be much more expensive; the heavy
concussions loosen the stratum of the roof and make it dangerous.

Large quantities of oil are used for lubrication and illumination.
"Zero" black oil and oils of that grade are used on the drills.
Miners' oil is generally used for illumination, though some of the
mines use a low grade of paraffine wax.

Two oil cans (each holding about 1J pints) are given to each
pair of drillers, one can for black oil and one for miners' oil.
These cans, properly filled, are given out to the men, as they go
on shift, at the "oil-house," located near the shaft underground.
This "oil-house" is in charge of the "oil boy," whose duty it is
to keep the cans clean, to fill them and to give them out at the
beginning of the shift. The cans are returned to the oil-house
at the end of the shift.

NOTES ON LEAD MINING 23

Kerosene is used in the hat-lamps in wet places.

The " oil-houses" are provided with three tanks. In some
instances these tanks are charged through pipes coming down
the shaft from the surface oil-house. These tanks are provided
with oil-pumps and graduated gage-glasses.

Shovelers or loaders operate in gangs of 8 to 12, and are
supervised by a "straw boss," who is provided with a gallon
can for illuminating oil. The cars are 20 cu. ft. ( 1 ton) capacity.
Under ordinary conditions one shoveler will load 20 of these cars
in a shift of 10 hours. They use " half -spring," long-handled,
round-pointed shovels.

Cars are of the solid-box pattern, and are dumped in cradles.
Loose and "Anaconda" manganese-steel wheels are the most
common. Gage of track, 24 to 30 in., 16 Ib. rails on main lines
and 12 Ib. on the side and temporary tracks. Cars are drawn
by mules. One mine has installed compressed-air locomotives
and is operating them with success.

Shafts are generally equipped with geared hoists, both steam
and electrically driven. Later hoists are all of the first-motion
pattern.

Generally the cars are hoisted to the top and dumped with
cradles. One shaft, however, is provided with a 5-ton skip,
charged at the bottom from a bin, into which the underground
cars are dumped. Upon arriving at the top the skip dumps
automatically. This design exhibits a number of advantages
over the older method and will probably find favor with other
mine operators.

THE LEAD ORES OF SOUTHWESTERN MISSOURI
BY C. V. PETRAEUS AND W. GEO. WARING

(October 21, 1905)

The lead ore of southwestern Missouri, and the adjoining area
in the vicinity of Galena, Kan., is obtained as a by-product of
zinc mining, the galena being separated from the blende in the
jigging process. Formerly the galena (together with " dry-bone,"
including cerussite and anglesite) was the principal ore mined
from surface deposits in clay, the blende being the subsidiary
product. In the deeper workings blende was found largely to
predominate; this is shown by the shipments of the district in
1904, which amounted to 267,297 tons of zinc concentrate and
34,533 tons of lead concentrate.

The lead occurs in segregated cubes, from less than one milli-
meter up to one foot in diameter. The cleavage is perfect, so
that each piece of ore when struck with a hammer breaks up
into smaller perfect cubes. In this respect the ore differs from
the galena encountered in the Rocky Mountain regions, where
torsional or shearing strains seem in most instances to have
destroyed the perfect cleavage of the minerals subsequent to
their original deposition. Cases of schistose and twisted structure
occur in lead deposits of the Joplin district but rarely, and they
are always quite local.

The separation of the galena from the blende and marcasite
("mundic") in the ordinary process of jigging is very complete;
the percentage of zinc and iron in the lead concentrate is insig-
nificant. As an illustration of this, the assays of 100 recent
consecutive shipments of lead ore from the district, taken at
random, are cited as follows:

7 shipments assayed from 57 to 70% lead
15 shipments assayed from 70 to 75% lead
46 shipments assayed from 75 to 79% lead
32 shipments assayed from 80 to 84.4% lead

Average of 100 shipments 78.4% lead

NOTES ON LEAD MINING 25

Fourteen shipment samples, ranging from 70 to 84.4 per cent,
lead, were tested for zinc and iron. These averaged 2.24 per cent.
Fe and 1.78 per cent. Zn, the highest zinc content being 4.5 per
cent. No bismuth or arsenic, and only very minute traces of
antimony, have ever been found in these ores. They contain only
about 0.0005 per cent, of silver (one-seventh of an ounce per ton)
and scarcely more than that of copper (occurring as chalcopyrite).

The pig lead produced from these ores is therefore very pure,
soft and uniform in quality, so that the term "soft Missouri
lead" has become a synonym for excellence in the manufacture
of lead alloys and products, such as litharge, red and white lead,
and orange mineral. Its freedom from bismuth, which is gener-
ally present in Colorado lead, makes it particularly suitable for
white lead; also for glass-maker's litharge and red lead. These
oxides, for use in making crystal glass, must be made by double
refining so as to remove even the small quantities of silver and
copper that are present. The resulting product, made from soft
Missouri lead, is far superior to any refined lead produced any-
where in this country or in Europe, even excelling the famous
Tarnowitz lead. It gives a luster and clarity to the glass that
no other lead will produce. Lead from southeastern Missouri,
Kentucky, Illinois, Iowa, and Wisconsin yields identical results,
but the refining is more difficult, not only because the lead con-
tains a little more silver and copper, but also because it contains
more antimony.

The valuation of the lead concentrate produced in the Joplin
district is based upon a wet assay, usually the molybdate or
ferrocyanide method. The price paid is determined variously.
One buyer pays a fixed price for average ore, making no deduc-
tions; as, for example, at present rates, $32.25 per 1000 Ib. whether
the ore assays 75 or 84 per cent. Pb, pig lead being worth $4.75
at St. Louis.1 Another pays $32.25 for 80 per cent, ore, or
over, deducting 50c. per unit for ores assaying under 80 per cent.
Another pays for 90 per cent, of the lead content of the ore as
shown by the assay, at the St. Louis price of pig lead, less a
smelting charge of, say, $6 to $8 per ton of ore.

The history of the development of lead ore buying in the
Joplin district is rather curious. In the early days of the district
the ore was smelted wholly on Scotch hearths, which, with the
1 The manuscript of this article was dated Oct. 5, 1905. ;

26 LEAD SMELTING AND REFINING

purest ores, would yield 70 per cent, metallic lead. No account
was taken of the lead in the rich slag, chemical determinations
being something unknown in the district at that time; it being
supposed generally that pure galena contained 700 Ib. lead to
the 1000 Ib. of ore, the value of 700 Ib. lead, less $4.50 per 1000 Ib.
of ore for freight and smelting costs, was returned to the miner.
The buyers graded the ore, according to their judgment, by its
appearance, as to its purity and also as to its behavior in smelting;
an ore, for example, from near the surface, imbedded in the
clay and coated more or less with sulphate, yielded its metal
more freely than the purer galenas from deeper workings.

This was the origin of the present method of buying — a
system that would hardly be tolerated except for the fact that
the lead is, as previously stated, considered a by-product of
zinc mining.

Originally all the lead ore from the Missouri-Kansas district
was smelted in the same region, either in the air furnace (rever-
beratory sweating-furnace) or in the water-back Scotch hearth.
Competition gradually developed in the market. Lead refiners
found the pure sulphide of special value in the production of
oxidized products. Some of the ore found its way to St. Louis,
and even as far away as Colorado, where it was used to collect
silver. Since the formation of the American Smelting and
Refining Company and the greatly increased output of the im-
mense deposits of lead ore in Idaho, no Missouri lead ore has
gone to Colorado.

Up to 1901, one concern had more or less the control of the
southwestern Missouri ores. At the present time, lead ore is
bought for smelters in Joplin, Carterville, and Granby, Mo.,
Galena, Kan., and Collinsville, 111., and complaint is heard that
present prices are really too high for the comfort of the smelters.
Yet the old principle of paying for lead ores upon the supposed
yield of 70 per cent., irrespective of the real lead content, is still
largely in vogue.

Any one interested in the matter will find it quite instructive
to calculate the returning charges, or gross profits, in smelting
these ores, on the basis of 70 per cent, recovery, at $32.25 per
1000 Ib. of ore, less 50c. per ton haulage, with lead at $4.77 per
100 Ib. at St. Louis. No deduction, it should be remarked, is
ever made for moisture in lead ores in this district. It is of

NOTES ON LEAD MINING 27

interest to observe that Dr. Isaac A. Hourwich estimates (ia the
U. S. Census Special Report on Mines and Quarries recently
issued) the average lead contents of the soft lead ores of Missouri
in 1902 at 68.2 per cent., taking as a basis the returns from five
leading mining and smelting companies of Missouri, which re-
ported a product of 70,491 tons of lead from 103,428 tons of
their own and purchased ore. The average prices for lead ore in
1902 were reported as follows, per 1000 lb.: Illinois, $19.53;
Iowa, $24.48; Kansas, $23.51; Missouri, $22.17; Wisconsin, $23.29;
Rocky Mountain and Atlantic Coast States, $10.90. In 1903,
according to Ingalls ("The Mineral Industry," Vol. XII), the ore
from the Joplin district commanded an average price of $53 per
2000 lb,, while the average in the southeastern district was $46.81.

PART II
ROAST-REACTION SMELTING

SCOTCH HEARTHS AND
REVERBERATORY FURNACES

LEAD SMELTING IN THE SCOTCH HEARTH

BY KENNETH W. M. MIDDLETON

(July 6, 1905)

In view of the fact that the Scotch hearth in its improved
form is now coming to the front again to some extent in lead
smelting, it may prove interesting to give a brief account of its
present use in the north of England.

Admitting that, where preliminary roasting is necessary,
the best results can be obtained with the water- jacketed blast
furnace (this being more especially the case where labor is an
expensive item), we have still as an alternative the method of
smelting raw in the Scotch hearth. At one works, which I
recently visited, all the ore was smelted raw; at another, all the ore
received a preliminary roast, and it is instructive to compare the
results obtained in the two cases. The following data refer to a
fairly "free-smelting" galena assaying nearly 80 per cent, of lead.

When smelting raw ore in the hearth, fully 7J long tons can
be treated in 24 hours, the amount of lead produced direct from
the furnace in the first fire being 8400 to 9000 Ib. ; this is equivalent
to 56 to 60 per cent, of lead, the remaining 24 to 20 per cent,
going into the fume and the slag.

When smelting ore which has received a preliminary roast of
two hours, 12,000 Ib. of lead is produced direct from the hearth,
this being equivalent to 65 per cent, of the ore. When the ore
is roasted, the output of the hearth is practically the same for
all ores of equal richness; but when smelting raw, if the galena
is finely divided, the output may fall much below that given
herewith; while, on the other hand, under the most favorable
conditions it may rise to 12,000 Ib. in 24 hours, or even more.

I had an opportunity of seeing a parcel of galena carrying
84 per cent, of lead (but broken down very fine) smelted raw.
The ore was kept damp and the blast fairly low; but, in spite of
that, a quantity of the ore was blown into the flue, and only
5100 Ib. of lead was produced from the hearth in 24 hours.

32 LEAD SMELTING AND REFINING

Galena carrying only 65 per cent, of lead does not give nearly
as satisfactory results when smelted raw in the hearth; barely
six tons of ore can be smelted in 24 hours, and only 4500 to
5400 Ib. of lead can be produced directly. This is equivalent to,
say, 43 per cent, of the ore in the first fire; the remaining 22 per
cent, goes into the slag or to the flue as fume. Moreover, the
65 per cent, ore requires 1500 Ib. of coal in 24 hours, while the
80 per cent, galena uses only 1000 Ib.

Turning now for a moment to the costs of smelting raw and
of smelting after a preliminary roast, we find that (in the case
of the two works we have been considering) the results are all in
favor of smelting raw, so far as a galena carrying nearly 80 per
cent, is concerned.

The cost of smelting, per ton of lead produced, is given here-
with:

ORE SMELTED RAW

Smelters' wages : $2.04

coal (425 Ib.) 0.38

Total $2.42

A very small quantity of lime is also used in this case for some ores, but
its cost would never amount to more than 4c. per ton of lead produced.

ORE RECEIVING A PRELIMINARY ROAST

Roasters' wages $0.61

" coal (425 Ib.) 0.65

Smelters' wages 1.08

coal (75 Ib.) 0.11

Peat and lime. . 0.08

Total $2.53

It should be noted also that the smelters at the works where
the ore was not roasted receive higher pay. In the eight-hour
shift they produce about 1J tons of lead; and as there are two
of them to a furnace, they make $3.06 between them, or $1.53
each. The two men smelting roasted ore produce about two
tons in an eight-hour shift, and therefore each receives $1.08
per shift.

Coming now to fume-smelting in the hearth, we can again
compare the results obtained in smelting raw and after roasting.
It is well to bear in mind, also, that, while only 6J per cent, of the
lead goes in the fume when smelting roasted ores in the hearth, a

ROAST-REACTION SMELTING 33

considerable larger proportion is thus lost when smelting raw ores.
When fume is smelted raw, it is best dealt with when containing
about 40 per cent, of moisture. One man attends to the hearth
(instead of two as when smelting ore), and in 24 hours 3000 Ib.
of lead is produced, the amount of coal used being 2100 Ib. No
lime is required.

When smelting roasted fume, two men attend to the hearth
and the output is 6000 Ib. in 24 hours, the amount of coal used
being 1800 Ib. In this latter case fluorspar happens to be avail-
able (practically free of cost), and a little of it is used with ad-
vantage in fume-smelting, as well as a small quantity of lime.

The cost of fume-smelting per ton of lead produced is given
herewith :

FUME SMELTED RAW

Smelters' wages $2.88

coal (1400 Ib.) 2.13

Total $5.01

FUME RECEIVING A PRELIMINARY ROAST

Roasters' wages $2.08

coal (1450 Ib.) : 2.18

Smelters' wages 2.04

coal (600 Ib.) 0.92

Peat and lime . . 0.08

Total $7.30

In this case, as in that of ore, the smelter of the raw fume
gets better pay; he has $1.44 per eight-hour shift, while the
smelter of the roasted ore has only $1.02 per eight-hour shift.

Fume takes four hours to roast, as compared to the two hours
taken by ore.

From these facts regarding Scotch-hearth smelting, it would
seem that with galena carrying, say, over 70 per cent, lead (but
more especially with ore up to 80 per cent, in lead, and, more-
over, fairly free from impurities detrimental to "free" smelting),
very satisfactory results can be obtained by smelting raw. Against
this, however, it must be said that at the works where the ore
is roasted attempts at smelting raw have been made several
times without sufficient success to justify the adoption of this
method, although the ores smelted average 75 per cent, lead and
seem quite suitable for the purpose.

LEAD SMELTING AND REFINING

Probably this may be accounted for by the fact that the
method of running the furnace when raw ore is being smelted
is rather different from that adopted when dealing with roasted
ore. Moreover, at the works under notice the furnaces are not
of the most modern construction; and, as the old custom of
dropping a peat in front of the blast every time the fire is made
up still survives, it is necessary to shut off the blast while this
is being done, and the fire is then apt to get rather slack.

The gray slag produced in the hearth is smelted in a small
blast furnace, a little poor fume, and sometimes a small quantity
of fluorspar, being added to facilitate the process. Some figures
regarding slag-smelting may be of interest. The slag-smelters
produce 9000 Ib. of lead in 24 hours. The cost of slag-smelting
per ton of lead produced is as follows:

Smelters' wages $1.60

Coke (1500 Ib.) 3.42

Peat... 0.06

Total.

$5.08

Recent analyses of Weardale (Durham county) lead smelted
in the Scotch hearth, and slag-lead smelted in the blast furnace,
are given herewith:

FUME-LEAD FROM
HEARTH

SILVER-LEAD FROM
HEARTH

SLAG-LEAD FROM
BLAST FURNACE

Lead

QQ Qf»7

QQ ni Q

Silver

fl ftO^s»

n n9nn

n 0149

Tin.

(1 OZ. 2 DWT. 21 GE.

PER LONG TON)

(6 OZ. 10 DWT. 16 OR.

PER LONG TON)
nil

(4 OZ. 12 DWT. 18 GR.

PER LONG TON)

Antimony

nil

•]

nil

n ft*7A

nil

U.o/4
n (Y)A.

Iron

OHIO

n ni o

U.UZ4
fi ft9Q

Zinc.

•i

UJUGM

•|

nil

99.9795

99.9960

99.9482

The ordinary form of the Scotch hearth is probably too well
known to need much description. The dimensions which have
been found most suitable are as follows: Front to back, 21 in.;
width, 27 in.; depth of hearth, 8 to 12 in. Formerly the distance
from front to back was 24 in., but this was found too much for
the blast and for the men.

The cast-iron hearth which holds the molten lead is set in

ROAST-REACTION SMELTING 35

brickwork; if 8 in. deep and capable of holding about f ton of
lead, it is quite large enough. The workstone or inclined plate
in front of the hearth is cast in one piece with it, and has a raised
holder on either side at the lower edge, and a gutter to convey
the overflowing lead to the melting-pot. The latter is best made
with a partition and an opening at the bottom through which
clean lead can run, so that it can be ladled into molds without
the necessity for skimming the dross off the surface. It is well
also to have a small fireplace below the melting-pot.

On each side of the hearth, and resting on it, is a heavy cast-
iron block, 9 in. thick, 15 in. high, 27 to 28 in. long. To save
metal, these are now cast hollow and air is caused to pass through
them. On the back of the hearth stands another cast-iron block
known as the "pipestone," through which the blast comes into
the furnace. In the older forms of pipestone the blast comes in
through a simple round or oval pipe, a common size being 3 or
4 in. wide by 2J in. high, and the pipestone is not water-cooled.
With this construction the hearth will not run satisfactorily
unless the pipestone is set with the greatest care, so as to have
the tuyere exactly in the center, and as there is no water-cooling
the metal quickly burns away when fume is being smelted.
Moreover, the blast is apt to be stopped by slag adhering to the
-end of the pipe. As already mentioned, a peat is dropped in
front of the blast every time the fire is made up, with the object
of keeping a clear passage open for the blast. This old custom
has, however, several serious disadvantages; first, it prevents the
blast being kept on continuously; and, second, it makes it neces-
sary to have the hearth open at the top so that the smelter-man
can go in by the side of it. In this case the ore is fed from the
side by the smelter-man, who works under the large hood placed
above the furnace to carry away the fume. Even when he is
engaged in shoveling back the fire from the front and is not
underneath the hood, it is impossible to prevent some fume from
blowing out; and there is much more liability to lead-poisoning
than when the hearth is closed at the top by the chimney and
the smelter-men work from the front. The best arrangement is
to have the hearth entirely closed in by the chimney, except for
the opening at the front, and to have a small auxiliary flue above
the workstone leading direct to the open air to catch any fume
that may blow out past the shutter in front of the hearth.

36 LEAD SMELTING AND REFINING

In an improved form of pipestone, a pipe connected to the
blast-main fits into the semicircular opening at the back and
is driven tight against a ridge in the flat side of the opening.
Going through the pipestone, the arch becomes gradually flatter,
and the blast emerges into the hearth, about 2 in. above the level
of the molten lead, through an oblong slit 12 in. long by 1 in.
wide, with a ledge projecting 1J in. immediately above it. The
back and front are similar, so that when one side gets damaged
the pipestone can be turned back to front.

Water is conveyed in a 2J-in. iron pipe to the pipestone, and
after passing through it is led away from the other end to a
water-box, which stands beside the hearth and into which the
red-hot lumps of slag are thrown to safeguard the smelters from
the noxious fumes.

On the top of the pipestone rests an upper backstone, also of
cast iron; it extends somewhat higher than the blocks at the
sides. All this metal above the level of the lead is necessary
because the partially fused lumps which stick to it have to be
knocked off with a long bar, so that if fire-bricks were used in
place of cast iron they would soon be broken up and destroyed.

With a covered-in hearth, when the ore is charged from the
front, the following is the method adopted in smelting raw ore:
The charge floats on the molten lead in the hearth, and at short
intervals the two smelters running the furnace ease it up with
long bars, which they insert underneath in the lead. Any pieces
of slag adhering to the sides and pipestone are broken off. . After
easing up the fire, the lumps of partially reduced ore, mixed with
cinders and slag, are shoveled on to the back of the fire; the slag
is drawn out upon the workstone (any pieces of ore adhering to
it being broken off and returned to the hearth), and it is then
quenched in a water-box placed alongside the workstone. One
or two shovelfuls of coal, broken fairly small and generally kept
damp, are thrown on the fire, together with the necessary amount
of ore, which is also kept damp if in a fine state of division. It
is part of the duty of the two smelters to ladle out the lead from
the melting-pot into the molds. In smelting ore a fairly strong,
steady blast is required, and it is made to blow right through so
as to keep the front of the fire bright. A little lime is thrown on
the front of the fire when the slag gets too greasy.

When smelting raw fume one man attends to the furnace. It

ROAST-REACTION SMELTING 37

does not have to be made up nearly as frequently, the work
being easier for one man than smelting ore is for two. The
unreduced clinkers and slag are dealt with exactly as in smelting
ore; and coal is also, in this case, thrown on the back of the fire,
but the blast does not blow right through to the front. On the
contrary, the front of the fire is kept tamped up with fume,
which should be of the coherency of a thick mud. The blast is
not so strong as that necessary for ore. The idea is partially to
bake the fume before submitting it to the hottest part of the
furnace, or to the part where the blast is most strongly felt. It
is only when smelting fume that it is necessary to keep the pipe-
stone water-cooled.

To start a furnace takes from two to three hours. The hearth
is left full of lead, and this has to be melted before the hearth is
in normal working order. Drawing the fire takes about three-
quarters of an hour; the clinkers are taken off and kept for starting
the next run, and the sides and back of the hearth are cleaned
down.

THE FEDERAL SMELTING WORKS, NEAR ALTON, ILL.1

BY O. PUFAHL
(June 2, 1906)

The works of the Federal Lead Company, near Alton, 111.,
were erected in 1902. They have a connection with the Chicago,
Peoria & St. Louis Railway, by which they receive all their raw
materials, and by which all the lead produced is shipped.

The ore smelted is galena, with dolomitic gangue, and a small
quantity of pyrites (containing a little copper, nickel, and cobalt)
from southeastern Missouri, and consists chiefly of fine concen-
trates, containing 60 to 70 per cent. lead. In addition thereto a
small proportion of lump ore is also smelted.

A striking feature at these works is the excellent facility for
handling the materials. The bins for the ore, coke and coal are
made of concrete and steel and are filled from cars running on
tracks laid above them. For transporting the materials about
the works a narrow-gage railway with electric locomotives is
used.

The ores are smelted by the Scotch-hearth process. There
are 20 hearths arranged in a row in a building constructed wholly
of steel and stone. The sump (4x2x1 ft.) of each furnace
contains about one ton of lead. The furnaces are operated with
low-pressure blast from a main which passes along the whole
row. The blast enters the furnace from a wind chest at the back
through eight 1-in. iron pipes, 2 in. above the bath of lead. The
two sides and the rear wall are cooled by a cast-iron water jacket
of 1 in. internal width.

Two men work, In eight-hour shifts, at each of the furnaces,
receiving 4.75 and 4.25c. respectively for every 100 Ib. of lead
produced. The ore is weighed out and heaped up in front of the
furnaces; on the track near by the coke is wheeled up in a flat
iron car with two compartments. The furnacemen are chiefly

'Translated from Zeit. f. Berg.- Hutten- und Salinenwesen, LIU (1905,
p. 460).

ROAST-REACTION SMELTING 39

negroes. At the side of each furnace is a small stock of coal,
which is used chiefly for maintaining a small fire under the lead
kettle. Only small quantities of coal are added from time to
time during the smelting operation.

Over each furnace is placed an iron hood, through which the
fumes and gases escape. They pass first through a collecting
pipe, extending through the whole works, to a 1500-ft. dust flue,
measuring 10 x 10 ft., in internal cross-section. Near the middle
of this is placed a fan of 100,000 cu. ft. capacity per minute,
which forces the fumes and gases into the bag-house, where they
are filtered through 1500 sacks of loosely woven cotton cloth,
each 25 ft. long and 18 in. in diameter, and thence pass up a
150-ft. stack.

The dust recovered in the collecting flue is burnt, together
with the fume caught by the bags, the coal which it contains
furnishing the combustible. It burns smolderingly and frits
together somewhat. The product (chiefly lead sulphate) is then
smelted in a shaft furnace, together with the gray slag from the
hearth furnaces. The total extraction of lead is about 98 per
cent., i.e., the combined process of Scotch-hearth and blast-fur-
nace smelting yields 98 per cent, of the lead contained in the
crude ore.

The direct yield of lead from the Scotch hearths is about
70 per cent. They also produce gray slag, containing much lead,
which amounts to about 25 per cent, of the weight of the ore.
About equal proportions of lead pass into the slag and into the
flue dust. When working to the full capacity, with rich ore
(80 per cent, lead and more) the 20 furnaces can produce about
200 tons of lead in 24 hours. The coke consumption in the
hearth furnaces amounts to only 8 per cent, of the ore. The
lead from these furnaces is refined for 30 minutes to one hour
by steam in a cast-iron kettle of 35 tons capacity, and is cast
into bars either alone or mixed with lead from the shaft furnace.
The "Federal Brand" carries nearly 99.9 per cent, lead, 0.05 to
0.1 per cent, copper, and traces of nickel and cobalt.

The working up of the between products from the hearth-
furnaces is carried out as follows: Slag, burnt flue dust and roasted
matte from a previous run, together with a liberal proportion of
iron slag (from the iron works at Alton), are smelted in a 12-tuyere
blast furnace for work-lead and matte. The furnace is provided

40 LEAD SMELTING AND REFINING

with a lead well at the back. The matte and slag are tapped off
together at the front and flow through a number of slag pots for
separation. The shells which remain adhering to the walls of
the pots on pouring out the slag are returned to the furnace.
All the waste slag (containing about 0.5 per cent, lead) is dumped
down a ravine belonging to the territory of the smeltery.

The lead from the shaft furnace is liquated in a small rever-
beratory furnace, of which the hearth consists of two inclined
perforated iron plates. The residue is returned to the shaft
furnace, while the liquated lead flows directly to the refining
kettle, which is filled in the course of four hours. Here it is
steamed for about one hour and is then cast into bars through a
Steitz siphon, after skimming off the oxide. The matte is crushed
and roasted in a reverberatory furnace (60 ft. long).

The power plant comprises three Stirling boilers and two
250-h. p. compound engines, of which one is for reserve; also one
steam-driven dynamo, coupled direct to the engine, furnishing
the current for the entire plant, for the electric locomotives, etc.

The coke is obtained from Pennsylvania and costs about $4 a
ton, while the coal comes from near-by collieries and costs $1 per
ton.

In the well-equipped laboratory the lead in the ores and slags
is determined daily by Alexander's (molybdate) method, while
the silver content of the lead (a little over 1 oz. per ton) is esti-
mated only once a month in an average sample. When the
plant is in full operation it gives employment to 150 men. . Cases
of lead-poisoning are said to occur but rarely, and then only in
a mild form.

LEAD SMELTING AT TARNOWITZ

(September 23, 1905)

The account of the introduction of the Huntington-Heberlein
process at Tarnowitz, Prussia, published elsewhere in this issue,
is of peculiar interest inasmuch as it tells of the complete dis-
placement by the new process of one of the old processes of lead
smelting which had become classic in the art. The roast-reaction
process of lead smelting, especially as carried out in reverberatory
furnaces, has been for a long time decadent, even in Europe.
Tarnowitz was one of the places where it survived most vigor-
ously.

Outside of Europe, this process never found any generally
extensive application. It was tried in the Joplin district, and
elsewhere in Missouri, with Flintshire furnaces in the seventies.
Later it was employed with modified Flintshire and Tarnowitz
furnaces at Desloge, in the Flat River district of Missouri, where
the plant is still in operation, but on a reduced scale.

The roast-reaction process of smelting, as practised at Tarno-
witz, was characterized by a comparatively large charge, slow
roasting and low temperature, differing in these respects from
the Carinthian and Welsh processes. It was not aimed to
extract the maximum proportion of lead in the reverberatory
furnace itself, the residue therefrom, which inevitably is high in
lead, being subsequently smelted in the blast furnace. Ores too
low in lead to be suitable for the reverberatory smelting were
sintered in ordinary furnaces and smelted in the blast furnace
together with the residue from the other process. In both of
these processes the loss of lead was comparatively high. One of
the most obvious advantages of the Huntington-Heberlein process
is its ability to reduce the loss of lead. The result in that respect
at Tarnowitz is clearly stated by Mr. Biernbaum, whose paper
will surely attract a good deal of attention.1

1 This paper is published in pp. 148-166 of this book.
41

LEAD SMELTING IN REVERBERATORY FURNACES AT
DESLOGE, MO.

BY WALTER RENTON INGALLS

(December 16, 1905)

The roast-reaction method of lead smelting in reverberatory
furnaces never found any general employment in the United
States, although in connection with the rude air-furnaces it was
early introduced in Missouri. The more elaborate Flintshire fur-
naces were tried at Granby, in the Joplin district, but they were
displaced there by Scotch hearths. The most extensive installa-
tion of furnaces of the Flintshire type was made at Desloge, in
the Flat River district of southeastern Missouri. This continued
in full operation until 1903, when the major portion of the plant
was closed, it being found more economical to ship the ore else-
where for smelting. However, two furnaces have been kept in
use to work up surplus ore. As a matter of historic interest, it is
worth while to record the technical results at Desloge, which have
not previously been described in metallurgical literature.

The Desloge plant, which was situated close to the dressing
works connected with the mine, and was designed for the smelting
of its concentrate, comprised five furnaces. The furnaces were of
various constructions. The oldest of them was of the Flintshire
type, and had a hearth 10 ft. wide and 14 ft. long. The other
furnaces were a combination of the Flintshire and Tarnowitz
types. They were built originally like the newer furnaces at
Tarnowitz, Upper Silesia, with a rather large rectangular hearth
and a lead sump placed at one side of the hearth near the throat
end; but good results were not obtained from that construction,
wherefore the furnaces were rearranged with the sump at one
side, but in the middle of the furnace, as in the Flintshire form.
The rectangular shape of the Tarnowitz hearth was, however,
retained. Furnaces thus modified had hearths 11 ft. wide and 16
ft. long, except one which had a hearth 13 ft. wide.

The same quantity of ore was put through each of these fur-

ROAST-REACTION SMELTING 43

naces, the increase in hearth area being practically of no useful
effect, because of inability to attain the requisite temperature in
all parts of the larger hearths with the method of heating em-
ployed. The men objected especially to a furnace with hearth
13 ft. wide, which it was found difficult to keep in proper condi-
tion, and also difficult to handle efficiently. Even the width of
11 ft. was considered too great, and preference was expressed
for a 10-ft. width. In this connection, it may be noted that
the old furnaces at Tarnowitz were 11 ft. 9 in. long and 10 ft.
10 in. wide, while the new furnaces were 16 ft. long and 8 ft. 10 in.
wide (Hofman, " Metallurgy of Lead," fifth edition, p. 112). All
of these dimensions were exceeded at Desloge.

The Flintshire furnaces at Desloge had three working doors
per side; the others had four, but only three per side were used,
the doors nearest the throat end being kept closed because of
insufficient temperature in that part of the furnace. The furnace
with hearth 11x14 ft. had a grate area of 6.5x3 ft. = 19.5
sq. ft.; the 11 x 16 furnaces had grates 8 x 3 = 24 ft. sq. The
ratios of grate to hearth area were therefore approximately 1 : 8
and 1 : 7.3, respectively. (Compare with ratio of 1 : 10 at Tarno-
witz, and 1 : 6§ at Stiperstones.) The ash pits were open from
behind in the customary English fashion. The grate bars were
cast iron, 36 in. long. The bars were 1 in. thick at the top, with
f-in. spaces between them. The open spaces were 32 in. long,
including the rib in the middle. The bars were 4 in. deep at the
middle and 2 in. at the ends. The distance from the surface of
the grate bars to the fire-door varied in the different furnaces.
Some of those with hearths 11 x 16 ft. and grates 8 x 3 ft. had the
bars 6 in. below the fire-door; in others the bars were almost on
a level with the fire-door.

The furnaces were run with a comparatively thin bed of coal
on the grate, and combustion was very imperfect, the percentage
of unburned carbon in the ash being commonly high. This was
unavoidable with the method of firing employed and the inferior
character of the coal (southern Illinois). The excessive con-
sumption of coal was due largely, however, to the practice of
raking out the entire bed of coal at the beginning of the operation
of "firing down" (beginning the reaction period), when a fresh
fire was built with cordwood and large lumps of coal.

Each furnace had two flues at the throat, 16 x 18 in. in size,

44 LEAD SMELTING AND REFINING

each flue being provided with a separate damper. Each furnace
had an iron chimney approximately 55 ft. high, of which 13 ft.
was a brick pedestal (64 x 64 in.) and the remaining 42 ft. sheet
steel, guyed. The chimneys were 42 in. in diameter. The dis-
tance from the outside end of the furnace to the chimney was
approximately 6 ft., and there was consequently but little oppor-
tunity for flue dust to collect in the flue. About once a month,
however, the chimney was opened at the base and about two
wheelbarrows (say 600 Ib.) of flue dust, assaying about 50 per
cent, lead, was recovered per furnace.

The furnace house was a frame building 45 ft. wide, with
boarded sides and a corrugated-iron pitch roof, supported by
steel trusses. The furnaces were set in this house. side by side,
their longitudinal axes being at right angles to the longitudinal
axis of the building. The distance from the outside of the fire-box
end of the furnace to the side of the building was 10 ft. The coal
was unloaded from a railway track alongside of the building and
was wheeled to the furnace in barrows. Some of the furnaces
were placed 18 ft. apart; others 22 ft. apart. The men much
preferred the greater distance, which made their work easier, an
important consideration in this method of smelting.

The hight from the floor to the working door of the furnace
was approximately 36 in. The working doors were formed with
cast-iron frames, making openings 7x11 in. on the inside and
15 x 28 in. on the outside. On the side of the furnace opposite
the middle working door was placed a cast-iron hemispherical
pot, set partially below the floor-line. This pot was 16 in. deep
and 24 in. in diameter; the metal was J in. thick. The distance
from the top of the pot to the line of the working door was 31 in.;
from the top of the pot to the bottom of the tap-door was 7 in.
The tap-door was 4 in. wide and 9 in. high, opening through a
cast-iron plate 1J in. thick. Below the tap-door and on a line
with the upper rim of the pot was a tap-hole 3J in. in diameter.
The frames of the working doors had lugs in front, against which
the buckstaves bore, to hold the frames in position. All other
parts of the sides of the furnace, including the fire-box, were
cased with f-in. cast-iron plates, which were obviously too light,
being badly cracked.

The cost of a furnace when built in 1893 was approximately
$1400, not including the chimney; but with the increased cost of

ROAST-REACTION SMELTING 45

material the present expense would probably be about $2000.
Notwithstanding the light construction of the furnaces, repairs
were never a large item. Once a month a furnace was idle about
24 hours while the throat was being cleaned out, and every two
months some repairing, such as relining the fire-boxes, etc., was
required. If repairs had to be made on the inside of the furnace,
two days would be lost while it was cooling sufficiently for the
men to enter. In refiring a furnace, from 8 to 12 hours was
required to raise it to the proper temperature. Out of the 365
days of the year, a furnace would lose from 20 to 25 days, for
cleaning the throat and making repairs to the fire-box, arch, etc.
When a furnace was run with two shifts the schedule of
operation was as follows:

Drop charge 4 a. m.

Begin work 7 a. m.

Begin firing down 11 a. m.

Begin first tapping 1p.m.

Rake out slag 2.30 p. m.

Begin second tapping 3 p. m.

Drop charge 4p.m.

Begin working 5.30 p. m.

Begin firing down 11 p. m.

Begin first tapping 1 a. m.

Rake out slag 2.30 a. m.

Begin second tapping 3 p. m.

With three shifts on a furnace, the schedule was as follows:

Drop charge 7 a. m.

Begin firing down 12 a. m.

Begin tapping 1p.m.

Rake out slag 2p.m.

Begin tapping 2.30 p. m.

Drop charge 3p.m.

Begin firing down 8 p. m.

Begin tapping 9 p. m.

Rake out slag 10 p. m.

Begin tapping 10.30 p. m.

Drop charge 11.00 p. m

Begin firing down 4 a. m.

Begin tapping 5 a. m.

Rake out slag 6 a. m.

pn tapping 6.30 a. m.

The hearths were composed of about 8 in. of gray slag beaten

46 LEAD SMELTING AND REFINING

down solidly on a basin of brick, which rested on a filling of clay,
rammed solid. The hearth was patched if necessary after the
drawing of each charge.

The system of smelting was analogous to that which was
practiced in Wales rather than to the Silesian, the charges being
worked off quickly, and with the aim of making a high extraction
of lead directly and a gray slag of comparatively low content in
lead. The average furnace charge was 3500 Ib. At the beginning
of the reaction period about 85 to 100 Ib. of crushed fluorspar
was thrown into the furnace and mixed well with the charge.
The furnace doors were then closed tightly and the temperature
raised, the grate having previously been cleaned. At the first
tapping about 1200 Ib. of lead would be obtained. A small
quantity of chips and bark was thrown into the lead in the kettle,
which was then poled for a few minutes, skimmed, and ladled
into molds, the pigs weighing 80 Ib. The skimmings and dross
were put back into the furnace. The pig lead was sold as " ordi-
nary soft Missouri." The gray slag was raked out of the furnace,
at the end of the operation, into a barrow, by which it was wheeled
to a pile outside of the building. Shipments of the slag were
made to other smelters from time to time, 95 per cent, of its
lead content being paid for when its assay was over 40 per cent.,
and 90 per cent, when lower.

Each furnace was manned by one smelter ($1.75) and one
helper ($1.55) per shift, when two shifts per 24 hours were run.
They had to get their own coal, ore and flux, and wheel away
their gray slag and ashes. In winter, when three shifts were run,
the men were paid only $1.65 and $1.50 respectively. There was
a foreman on the day shift, but none at night. The total coal
consumption was ordinarily about 0.8 to 0.9 per ton of ore.
Run-of-mine coal was used, which cost about $2 per ton delivered.
The coal was of inferior quality, and it was wastefully burned,
as previously referred to, wherefore the consumption was high in
comparison with the average at Tarnowitz, where it used to be
about 0.5 per ton of ore.

The chief features of the practice at Desloge are compared
with those at Tarnowitz, Silesia and Holy well (Flintshire), and
Stiperstones (Shropshire), Wales, in the following table, the data
for Silesia and Wales being taken from Hofman's "Metallurgy of
Lead," fifth edition, pp. 112, 113.

ROAST-REACTION SMELTING

DETAIL

HOLYWELL

STIPER-

STONES

TARNOWITZ

DESLOGE

Hearth length ft

12.00

9.75

11.75

1600

16 00

9.50

10.83

8.83

1100

4.50

8.00

800

2.50

1.67

3 00

1:8

1:6*

1:10

1-10

1-71

3 *

3 ; *

Ore smelted per 24 hr., Ib

7,050

8,800

16,500

10,500

iU.OW

Assay of ore % Pb ....

75-80

77.5

70-74

Gray slag % of charge

Gray slag' % Pb *"

38.8

Coal used per ton ore

0.57-0.76

0.56

0.46

0.50

0.90

The regular furnace charge at Desloge was 3500 Ib. The
working of three charges per 24 hours gave a daily capacity of
10,500 Ib. per furnace. These figures refer to the wet weight
of the concentrate, which was smelted just as delivered from
the mill. Its size was 9 mm. and finer. Assuming its average
moisture content to be 5 per cent., the daily capacity per furnace
was about 10,000 Ib. (5 tons) of dry ore.

The metallurgical result is indicated by the figures for two
months of operation in 1900. The quantity of ore smelted was
1012 tons, equivalent to approximately 962 tons dry weight.
The pig lead produced was 523.3 tons, or 54.4 per cent, of the
weight of the ore. The gray slag produced was 262.25 tons, or
about 27 per cent, of the weight of the ore. The assay of the
ore was approximately 70 per cent, lead, giving a content of
673.4 tons in the ore smelted. The gray slag assayed approxi-
mately 38 per cent, lead, giving a content of 99.66 tons. As-
suming that 90 per cent, of the lead in the gray slag be recoverable
in the subsequent smelting in the blast furnace, or 89.7 tons,
the total extraction of lead in the process was 523.3 + 89.7 ~-
673.4 = 91 per cent. The metallurgical efficiency of the process
was, therefore, reasonably high, especially in view of the absence
of dust chambers.

The cost of smelting with five furnaces in operation, each
treating three charges per day, was approximately as follows:

1 foreman at $3 $3.00

5 furnace crews at $9.90 49.60

Unloading 21 tons of coal at 6c 1.26

Loading 14 tons lead at 15c 2.10

" 7 tons gray slag at 15c 1.05

Total labor.. . $56.91

48 LEAD SMELTING AND REFINING

21 tons coal at $2 $42.00

Flux and supplies 13.00

Blacksmithing and repairs 10.00

Total .$121.91

On the basis of 6.25 tons of wet ore, this would be $4.65 per
ton. The actual cost in seven consecutive months of 1900 was
as follows: Labor, $1.98 per ton; coal, $1.86; flux and supplies,
$0.51; blacksmithing and repairs, $0.39; miscellaneous, $0.017;
total, $4.757. If the cost of smelting the gray slag be reckoned
at $8 per ton, and the proportion of gray slag be reckoned at
0.25 ton per ton of galena concentrate, the total cost of treatment
of the latter comes to about $6.75 per ton of wet charge, or about
$7 per ton of dry charge. This cost could be materially reduced
in a larger and more perfectly designed plant.

The practice at Desloge did not compare unfavorably, either
in respect to metal extracted or in smelting cost, with the roast-
reduction method of smelting or the Scotch hearth method, as
carried out in the plants of similar capacity and approximately
the same date of construction, smelting the same class of ore,
but the larger and more recent plants in the vicinity of St. Louis
could offer sufficiently better terms to make it advisable to close
down the Desloge plant and ship the ore to them. One of the
drawbacks of the reverberatory method of smelting was the
necessity of shipping away the gray slag, the quantity of that
product made in a small plant being insufficient to warrant the
operation of an independent shaft furnace.

PART III
SINTERING AND BRIQUETTING

THE DESULPHURIZATION OF SLIMES BY HEAP
ROASTING AT BROKEN HILL1

BY E. J. HORWOOD

(August 22, 1903)

It is well known that, owing to the intimate mixture of the
constituents of the Broken Hill sulphide ores, a great deal of
crushing and grinding is required to detach the particles of galena
from the zinc blende and the gangue; and it will be understood,
therefore, that a considerable amount of the material is converted
into a slime which consists of minute but well-defined particles
of all the constituents of the ore, the relative proportions of
which depend on the dual characteristics of hardness and abun-
dance of the various constituents. An analysis of the slime shows
the contents to be as follows:

Galena (PbS) 24.00

Blende (ZnS) 29.00

Pyrite (FeS2) 3.38

Ferric oxide (Fe2O3) 4.17

Ferrous oxide (FeO) contained in garnets 1.03

Oxide of manganese (MnO) contained in rhodonite and

garnets 6.66

Alumina (A12O3) contained in kaolin and garnets 5.40

Lime (CaO) contained in garnets, etc 3.40

Silica (SiO2) 22.98

Silver (Ag) .06

100.48

Galena, being the softest of these, is found in the slimes to a
larger extent than in the crude ore; it is also, for the same reason,
in the finest state of subdivision, as is well illustrated by the
fact that the last slime to settle in water is invariably much the
richest in lead, while the percentages of the harder constituents,
zinc blende and gangue, show a corresponding reduction in

1 Abstract from Transactions of the Australasian Institute of Mining
Engineers, Vol. IX, Part 1.

52 LEAD SMELTING AND REFINING

quantity, by reason of their being generally in larger sized particles
and consequently settling earlier.

The fairly complete liberation of each of the constituent
minerals of the ore that takes place in sliming tends, of course,
to help the production of a high-grade concentrate by the use of
tables and vanners, and undoubtedly a fair recovery of lead is
quite possible, even with existing machines, in the treatment of
fine slimes; but, owing to the great reduction in the capacity
of the machines, which takes place when it is attempted to carry
the vanning of the finer slimes too far, and the consequently
greatly increased area of the machines that would be necessary,
the operation, sooner or later, becomes unprofitable.

The extent to which the vanner treatment of slimes should
be carried is, of course, less in the case of those mines owning
smelters than with those which have to depend on the sale of
concentrates as their sole source of profit. In the case of the
Proprietary Company, all slime produced in crushing is passed
over the machines after classification. A high recovery of lead
in the form of concentrates is, of course, neither expected nor
obtained, for reasons already explained; but the finest lead-bearing
slimes are allowed to unite with the tailings, which are collected
from groups of machines, and are then run into pointed boxes,
where, with the aid of hydraulic classification, the fine rich slimes
are washed out and carried to settling bins and tanks, where the
water is stilled and allowed to deposit its slime, and pass over a
wide overflow as clear water. The slime thus recovered amounts
to over 1200 tons weekly, or about 11 per cent., by weight, of
the ore, and assays about 20 per cent, lead, 17 per cent, zinc,
and 18 oz. silver, and represents, in lead value, about 11 per
cent, of the original lead contents of the crude ore and rather
more than that percentage in silver contents. These slimes are
thus a by-product of the mills, and their production is unavoid-
able; but as they are not chargeable with the cost of milling, they
are an asset of considerable value, more especially so since it has
been demonstrated that they can be desulphurized sufficiently
for smelting purposes by a simple operation, and, at the same
time, converted into such a physical condition as renders the
material well suited for smelting, owing to its ability to resist
pressure in the furnaces.

The Broken Hill Proprietary Company has many thousands

SINTERING AND BRIQUETTING 53

of tons of these slimes which the smelters have hitherto been
unable to cope with, owing to the roasters being fully occupied
with the more valuable concentrates. Moreover, the desulphu-
rization of slimes in Ropp mechanical roasters is objectionable
for various reasons, namely, owing to the large amount of dust
created with such fine material, resulting injuriously to the men
employed; also on account of the reduction in the capacity of
the roasters, and consequent increase in working cost, owing to
the lightness of the slime, especially when hot, as compared with
concentrates, and the necessity for limiting the thickness of
material on the bed of the roasters to a certain small maximum.
Further, the desulphurization of the slimes is no more complete
with the mechanical roasters than in the case of heap roasting,
and the combined cost of roasting and briquetting being quite
three shillings (or 75c.) per ton in excess of the cost of heap
roasting, the latter possesses many advantages. These heaps are
being dealt with, preparatory to roasting, by picking down the
material in lumps of about 5 in. in thickness, while the fine dry
smalls, unavoidably produced, are worked up in a pug mill with
water, and dealt with in the same way as the wet slime produced
from current work.

The slime, as produced by the mills, is run from bins into
railway trucks in a semi-fluid condition, and shortly after being
tipped alongside one of the various sidings on the mine is in a
fit condition to be cut with shovels into rough bricks, which dry
with fair rapidity, and when required for roasting are easily re-
loaded into railway trucks. As each man can cut about 20 tons
of bricks per day, the cost is small. Various other methods of
lumping the slime were tried, including trucking the semi-fluid
material on movable trams, alongside which were set laths, about
9 in. apart, which enabled long slabs to be formed 9 in. wide
and 5 in. thick, which were, after drying, picked up in suitable
lumps and loaded in platform trucks, thence on railway trucks.
Owing to the inferior roasting that takes place with bricks having
flat sides, which are liable to come into close contact in roasting,
and to the rather high labor cost, this method was discontinued.
Another method was to allow the slime to dry partially after
being emptied from railway trucks, and to break it into lumps
by means of picks; but this method entailed the making of an
increased amount of smalls, besides taking up more siding room,

54 LEAD SMELTING AND REFINING

owing to the extra time required for drying, as compared with
the method now in use. Ordinary bricking machines could, of
course, be used, but when the cost of handling the slime before
and after bricking is counted, the cost would be greater than
with the simple method now in use; the material being in too
fluid a condition for making into bricks until some time elapses
for drying, a double handling would be necessitated before sending
it to the bricking machine. If, however, the slime could be allowed
time to dry sufficiently in the trucks, bricking by machinery would
probably be preferable. Rather more than 10 per cent, of smalls
is made in handling the lumps in and out of the railway trucks,
and this is, as already noted, worked up with water in a pug
mill at the sintering works, and used partly for covering the
heaps with slime to exclude an excessive amount of air. The
balance is thrown out and cut into bricks, as already described.

At the heaps the lumps are at present being thrown from
one man to another to reach their destination in the heap, but
the sidings have been laid out in duplicate with a view to enabling
traveling cranes to be used on the line next the heap, the lumps
to be loaded primarily into wooden skips fitting the trucks. It is
probable, however, that the lumps will require to be handled
out of the skips into their place in the heap, as the brittle nature
of the material may be found to render automatic tipping im-
practicable. A considerable saving in labor would nevertheless
accompany the use of cranes, which would likewise be advan-
tageous in loading the sintered material.

In order to reduce the inconvenience arising from fumes,
length is very desirable in siding accommodation, so that heap
building may be carried on at a sufficient distance from the
burning kilns. It is for the same reason preferable to build in a
large tonnage at one time, lighting the heaps altogether. As
the heaps burn about two weeks only, long intervals intervene,
during which the fumes are absent.

In the experimental stages of slime roasting, fuel, chiefly
wood, was used in quantities up to 5 per cent., and was placed
on the ground at the bottom of the heap, where also a number
of flues, loosely built bricks, were placed for the circulation of air.
The amount of fuel used has, however, been gradually reduced,
until the present practice of placing no fuel whatever in the
bottom was arrived at; but instead less than 1 per cent, of wood

SINTERING AND BRIQUETTING 55

is now burned in small enlargements of the flues, under the outer
portion of the pile, and placed about 12 ft. apart at the centers.
This is found to be sufficient to start the roasting operation within
24 hours of lighting, after which no further fuel is necessary.

As regards the dimensions of the heaps, the width found
most suitable is 22 ft. at the base, the sides sloping up rather
flatter than one to one, with a flat section on top reaching about
7 ft. in hight. As there is always about 6 in. of the outer crust
imperfectly roasted, it is advisable to make the length as great
as possible, thus minimizing the surface exposed. The company
is building heaps up to 2000 ft. long.

During roasting care is required to regulate the air supply,
the object being to avoid too fierce a roast, which tends to sinter
and partially fuse the material on the outer portions of the lumps,
while inside there is raw slime. By extending the roast over a
longer period this is avoided, and a more complete desulphuriza-
tion is effected. Experiments conducted by Mr. Bradford, the
chief assayer, demonstrated that, at a temperature of 400 deg. C.,
the sulphide slime is converted into basic sulphate, while at a
temperature of 800 deg. C. the material becomes sintered owing
to the decomposition of the basic sulphate and the formation of
fusible silicate of lead.

In practice, the sulphur contents of the material, which
originally are about 14 per cent., become reduced to from 6.5 to
8.5 per cent., half in the form of basic sulphate and half as sul-
phides; much of the material sinters and becomes matted together
in a fairly solid mass. The heaps are built without chimneys of
any kind; a strip about 5 ft. wide along the crest of the pile is
left uncovered by plastered slime, and this, together with the
open way in which the lumps are built in, allows a natural draft
to be set up, which can be regulated by partly closing the open
ends of the flues at the base of the pile. Masonry kilns were
used in the earlier stages with good results, which, however, were
not so much better than those obtained by the heap method as
to justify the expense of building, taking into consideration, too,
the extra cost of handling the roasted material in the necessarily
more confined space.

Much interest has been taken in the chemical reactions which
take place in the operation of desulphurization of these slimes, it
being contended, on the one hand, that the unexpectedly rapid

56 LEAD SMELTING AND REFINING

roast which takes place may be due to the sulphide being in a
very fine state of subdivision, and more or less porous, thus
allowing the air ready access to the sulphur, producing sulphurous
acid gas (SO2). On the other hand, others, of whom Mr. Car-
michael is the chief exponent, claim that several reactions take
place during the operation, connected with the rhodonite and
lime compounds present in the slimes, which he describes as
follows :

"The temperature of the kilns having reached a dull, red
heat, the rhodonite (silicate of manganese) is converted into
manganous oxide and silica; at a rather higher temperature the
calcium compounds are also split up, with formation of calcium
sulphide, the sulphur being provided by the slimes. The air
permeating the mass oxidizes the manganese oxide and calcium
sulphide into manganese tetroxide and calcium sulphate respec-
tively, as shown as follows:

3MnO + O = Mn3O4
CaS + 4O = CaSO4,

and, as such, are carriers of a form of concentrated oxygen to
the sulphide slimes, with a corresponding reduction to manga-
nous oxide and calcium sulphide, as shown by the following
equation, in the case of lead:

PbS + 4Mn3O4 = PbSO4 + 12MnO
PbS + CaSO4 = PbSO4 + CaS.

The oxidation of the manganous oxide and calcium sulphide is
repeated, and these alternate reactions recur until the desul-
phurization ceases, or the kiln cools down to a temperature below
which oxidation cannot occur. These reactions, being heat-pro-
ducing, provide part of the heat necessary for desulphurization,
which is brought about by certain concurrent reactions between
metallic sulphates and sulphide.

"The first that probably occurs is that in which two equiva-
lents of the metallic sulphide react on one of the metallic sulphate
with reduction to the metal, metallic sulphide, and sulphurous
acid, as shown by the following equation in the case of lead:

2PbS + PbSO4 = 2Pb + PbS + 2SO2.

SINTERING AND BRIQUETTING 57

"The metal so formed, in the presence of air, is oxidized, and
in this state reacts on a further portion of the metallic sulphide
produced, with an increased formation of metal and evolution of
sulphurous acid, according to the following equation, in the case
of lead:

2PbO + PbS = Pb + S02.

" The metal so produced in this reaction is wholly reoxidized
by the oxygen of the air current, and being free to react on still
further portions of the metallic sulphide, repeats the reaction,
and becomes an important factor in the desulphurizing of the
undecomposed portion of the material. As the desulphurization
proceeds, and the sulphate of metal accumulates, reactions are
set up between the metallic sulphide and different multiple pro-
portions of the metallic sulphate, with the formation of metal,
metallic oxide, and evolution of sulphurous acid, as follows:

"With two equivalents of metallic sulphate to one equivalent
of metallic sulphide, in the case of lead, according to the following
equation:

PbS + 2PbSO4 = 2PbO + Pb + 3SO2.

" With three equivalents of metallic sulphate to one of metallic
sulphide, in the case of lead, according to the following equation:

PbS + 3PbSO4 = 4PbO + 4SO2."

The volatility of sulphide of lead — especially in the presence
of an inert gas such as sulphurous acid — being greater than that
of the sulphate, oxide, or the metal itself, it might be thought
that the conditions are conducive to a serious loss of lead. This,
however, is reduced to a minimum, owing to the easily volatilized
sulphide being trapped, as non- volatile sulphate, by small portions
of sulphuric anhydride (SO3), which is formed by a catalytic
reaction set up between the hot ore, sulphurous acid, and the air
passing through the mass. Owing to the non-volatility of the
silver compounds in the slimes, the loss of this metal has been
found to be inappreciable. The zinc contents of the slime are
reduced appreciably, thus rendering the material more suitable
for smelting. After desulphurization ceases, a few days are
allowed for cooling off. On the breaking up of the mass for

58 LEAD SMELTING AND REFINING

despatch to the smelters, as much of the lower portion of the
walls is left intact as possible, so that it can be utilized for
the next roast, thus avoiding the re-building of the whole of
the walls.1

1In the course of subsequent discussion Mr. Horwood stated that the
losses in roasting were 12J per cent, in lead and probably about 5 per cent.
in silver. As compared to roasting in Ropp furnaces the loss in lead was
5 to 6 per cent, greater, but the difference of loss in silver was, he thought,
not appreciable. Mr. Hibbard said that the Central mine had obtained
similar satisfactory results with masonry kilns. — EDITOR.

THE PREPARATION OF FINE MATERIAL FOR
SMELTING

BY T. J. GREEN WAY

(January 12, 1905)

In the course of smelting, at the works of the company known
as the Broken Hill Proprietary Block 14, material which consisted
chiefly of silver-lead concentrate and slime, resulting from the
concentration of the Broken Hill complex sulphide ore, I had to
contend with all the troubles which attend the treatment of large
quantities of finely divided material in blast furnaces. With the
view of avoiding these troubles, I experimented with various
briquetting processes; and, after a number of more or less unsat-
isfactory experiences, I adopted a procedure similar to that
followed in manufacturing ordinary bricks by what is known as
the semi-dry brick-pressing process. This method of briquetting
not only converts the finely divided material cheaply and effect-
ively into hard semi-fused lumps, which are especially suitable
for the heavy furnace burdens required by modern smelting
practice, but also eliminates sulphur, arsenic, etc., to a great ex-
tent; therefore, it is capable of wide application in dealing with
concentrate, slime, and other finely divided material containing
lead, copper and the precious metals.

This briquetting process comprises the following series of
operations:

1. Mixing the finely divided material with water and newly
slaked lime.

2. Pressing the mixture into blocks of the size and shape of
ordinary bricks.

3. Stacking the briquettes in suitably covered kilns.

4. Burning the briquettes, so as to harden them, without
melting, at the same time eliminating sulphur, arsenic, etc.

1. The material is dumped into a mixing plant, together with
such proportions of screened slacked lime (usually from three to
five per cent.) and water as shall produce a powdery mixture,

60 LEAD SMELTING AND REFI.MXd

which will, on being squeezed in the hand, cohere into dry lumps.
In preparing the mixture, it is well to mix sandy material with
suitable proportions of fine, such as slime, in order that the finer
material may act as a binding agent.

The mixer used by me consists of an iron trough, about 8 ft.
long, traversed by a pair of revolving shafts, carrying a series of
knives arranged screw-fashion; and so placed that the knives on
one shaft travel through the spaces between the knives on the
other shaft. The various materials are dumped into one end of
the mixing trough, from barrows or trucks, and are delivered
continuously at the other end of the trough, into an elevator
which conveys the mixture to the brick-pressing plant.

2. The plant employed was the semi-dry brick-press. This
machine receives the mixture from the elevators, and delivers it
in the form of briquettes, which can at once be stacked in the
kilns. It was found that such material as concentrate and slime
has comparatively little mobility in the dies during the pressing
operation; this necessitates the use of a device which provides
for the accurate filling of the dies. It was also found that the
materials treated by smelters vary in compressibility, and this
renders necessary the adoption of a brick-pressing plant having
plungers which are forced into the dies by means of adjustable
springs, brick-presses having plungers actuated by rigid mecha-
nism being extremely liable to jam and break.

3. Briquettes made from such material as concentrate and
slime vary in fusibility; they are also combustible, and while
being burned they produce large quantities of smoke containing
sulphurous acid and other objectionable fumes. It is therefore
necessary that s.uch briquettes be burned in kilns provided with
arrangements for accurately controlling the burning operations,
and for conveniently disposing of the smoke. Suitable kilns,
which will contain from 30 to 50 tons of briquettes per setting,
are employed for this purpose. Regenerative kilns of the Hoff-
man type might be used for dealing with some classes of material,
but, for general purposes, the kilns as designed here will be found
more convenient.

The briquettes are stacked according to the character of the
material and the object to be obtained. The various methods
of stacking, and the reasons for adopting them, can be readily
learned by studying ordinary brick-burning operations in any

SINTERING AND BRIQUETTING 61

large brick-yard. After the stacking is complete the kiln-fronts
are built up with burnt briquettes produced in conducting previous,
operations, and all the joints are well luted.

4. In burning briquettes made from pyrite or other self-
burning material, it is simply necessary to maintain a fire in the
kiln fireplaces for a period of from 10 to 20 hours. When it is
judged that this firing has been continued long enough, the
fire-bars are drawn and the fronts are luted with burnt briquettes
in the same manner as the kiln-fronts. Holes about two inches
square are then made in these lutings, through which the air
required for the further burning of the briquettes is allowed to
enter the kilns under proper control. After the fireplaces are
thus closed the progress of the burning, which continues for
periods of from three to six days, is watched through small in-
spection holes made in the kiln-fronts; and when it is seen that
the burning is complete the fronts are partially torn away, in
order to accelerate the cooling of the burnt briquettes, which
are broken down and conveyed to the smelters as soon as they
can be conveniently handled.

When briquettes made from pyrite concentrate, or of other
free-burning material, are thus treated, they are not only sintered
but they are also more or less effectively roasted, and it may be
taken for granted that any ore which can be effectively roasted
in the lump form in kilns or stalls will form briquettes that will
both sinter and roast well; indeed, one may say more than this,,
for briquettes which will sinter and roast well can be made from
many classes of ore that cannot be effectively treated by ordinary
kiln- and stall-roasting operations; and, moreover, good-burning
briquettes may be made from mixtures of free-burning and poor-
burning material. Briquettes containing large proportions of
pyrite or other free-burning material will, unless the air-supply
is properly controlled, often heat up to such an extent as to fuse
into solid masses, much in the same manner as matte of pyritie
ore will melt when it is unskilfully handled in roasting. In
dealing with material which will not burn freely, such as roasted
concentrate, the briquetting is conducted with the intention of
sintering the material; and in this case the firing of the kilns is
continued for periods of from three to four days, the procedure
being similar in every way to that followed in burning ordinary
bricks.

62 LEAD SMELTING AND REFINING

When conducting my earlier briquetting operations I made
the briquettes by simply pugging the finely divided material,
following a practice similar to that adopted in producing " slop-
made" bricks by hand. This method of making the briquettes
was attended with a number of obvious disadvantages, and was
abandoned as soon as the semi-dry brick-pressing plant became
available. The extent to which this process, or modifications of
it, may be applied is shown by the fact that, following upon infor-
mation given by me, the Broken Hill Proprietary Company
adopted a similar method of sintering and roasting slime, con-
sisting of about 20 per cent, galena, 20 per cent, blende, and 60
per cent, silicious gangue. The procedure followed in this case
consisted of simply pugging the slime, and running the pug upon
a floor to dry; afterward cutting the dried material into lumps
by means of suitable cutting tools, and then piling the lumps
over firing foundations, following a practice similar to that pur-
sued in conducting ordinary heap-roasting. This company is
now treating from 500 to 1000 tons of slime weekly in this manner.
It is, however, certain that better results would attend the treat-
ment of this material by making this slime into briquettes and
burning them in kilns.

The cost of briquetting and burning material in the manner
first described, with labor at 25c. per hour, and wood or coal at
$4 per ton, amounts to from $1 to $1.50 per ton of material.

THE BRIQUETTING OF MINERALS

BY ROBERT SCHORR

(November 22, 1902)

The value of briquetting in connection with metallurgical
processes and the manufacture of artificial stone is well understood
and appreciated. In smelting plants there is always more or less
flue dust, fine ores, and sometimes fine concentrates to be treated,
but the charging of such fine material directly into a furnace
would cause trouble and irregularities, and would lessen its
capacity also. As mineral briquetting cannot be effected without
considerable wear upon the machinery and without quite appre-
ciable expense in binder, labor, and handling, many smelters try
to avoid it.

The financial question, however, is not as serious as it may at
first appear, and taking the large output of modern briquetting
machines in consideration, the cost for repairs amounts only to
a few cents per ton of briquetted material. The total cost depends
in the first place on the cost of labor, power and the binder, and
in most American smelters it varies between $0.65 and $1.25
per ton of briquettes.

Ordinary brick presses, with clay as a binder, were used in
Europe as well as in this country, but they are too slow and
expensive for large propositions and the presence of clay is usually
undesirable.

The English Yeadon (fuel) press has also been used for some
years at the Carlton Iron Company's Works at Ferryhill in
England, and at the Ore and Fuel Company's plant at Coatbridge
in the same country; also by some Continental firms. Dupuis &
Sons, Paris, furnished a few presses which are mostly used for
manganese and iron ores and pyrites. In some localities coke dust
is added. The making of clay briquettes or mud-cakes is the
crudest form of briquetting; but while heat has to be expended
to evaporate the 40 to 50 per cent, of moisture in them, and while
considerable flue dust is made, this method is better than feeding
fine ore or flue dust directly into the furnace.

64 LEAD SMELTING AND REFINING

The only other method of avoiding briquetting is by fusing
ore fines in slagging reverberatory furnaces and by adding flue
dust in the slagging pit, thus incorporating it with the slagging
ore. This is practised sometimes in silver-lead smelters, but in
connection with copper or iron smelters it is not practicable.

In briquetting minerals a thorough mixing and kneading is
of the first importance. If this is done properly a comparatively
low pressure will suffice to create a good and solid briquette,
which after six to eight hours of air-drying, or after a speedier
elimination of the surplus of moisture in hot-air chambers, will
be ready for the furnace charge. A good briquette should permit
transportation without excessive breakage or dust a few hours
after being made, and it should retain its shape in the furnace
until completely fused, so as to create as little flue dust as possible.
The briquette should be dense, otherwise it will crumble under
the influence of bad weather.

The two presses on the American machinery market are the
type built by the Chisholm, Boyd & White Company, of Chicago,
and the briquetting machine manufactured by the H. S. Mould
Company, of Pittsburg. Both are extensively used, and in many
metallurgical plants it will pay well to adopt them.

From 4 to 6 per cent, of milk of lime is generally used as
binder, and this has a desirable fluxing influence also. A com-
plete outfit comprises, besides the press, a mixer for slacking the
lime, and a feed-pump which discharges the liquid in proportion
into the main mixer wherein the ore fines, flue dust, or concen-
trates are shoveled.

The Chisholm, Boyd & White Company's press makes 80
briquettes per minute, which, with a new disk, are of 4 in. diam-
eter and 2J in. hight, thus giving about 872 cu. ft. of briquette
volume per 10 hours, or 50 to 80 tons, depending on the weight
of the material. With the wear of the disk the hight of the
briquettes is reduced and consequently the capacity of the machine
also. The disk weighs about 1600 lb., and as most large smelters
have their own foundries it can be replaced with little expense.
About 30 effective horse-power is usually provided for driving
the apparatus. The machine is too well known to metallurgists
and engineers to require further comment or description.

The H. S. Mould Company has also succeeded in making its
machine a thorough practical success. This machine is a plunger-

SINTERING AND BRIQUETTING 65

type press. The largest press built employs six plungers, and at
25 revolutions it makes 150 briquettes of 3 in. diameter and 3 in.
hight, or 1080 cu. ft. per 10 hours. Its rated capacity is 100
tons per 10 hours.

In using a plunger-type press the material should not contain
more than 7 per cent, mechanical moisture. If wet concentrates
have to be briquetted it is necessary to add dry ore fines or flue
dust to arrive at a proper consistency. The briquettes are very
solid and only air-drying for a few hours is necessary.

The cylindrical shape of briquettes is very good, as it insures
a proper air circulation in the furnace and consequently a rapid
oxidation and fusion.

The wear of the Mould Company's press is mostly confined
to the chilled iron bushings and to the pistons. Auxiliary ma-
chinery consists of the slacker, the feeder and the main mixer.
The press is of a very substantial design, and it is claimed that
the cost of repairs does not amount to more than 3c. per ton
of briquettes.

Wear and tear is unavoidable in a crude operation like bri-
quetting; to treat flue dust, ore fines, and fine concentrates
successfully, it is almost absolutely necessary to resort to it.

Edison used a number of intermittent-acting presses at his
magnetic iron-separation works in New Jersey, but this plant
shut down some time ago.

A BRICKING PLANT FOR FLUE DUST AND FINE ORES

BY JAMES C. BENNETT

(September 15, 1904)

The plant, which is here described, for bricking fine ores and
flue dust, was designed and the plans produced in the engineering
department of the Selby smelter. The machinery contained in
the plant consists of a Boyd four-mold brick press, a 7-ft. wet
pan or Chile mill, a 50-h.p. induction motor, and a conveyor-
elevator, together with the necessary pulleys and shafting.

The press, Chile mill, and motor need no special mention, as
they all are from standard patterns and bought, without altera-
tions, from the respective builders. The Chile mill was purchased
from the builders of the brick press. The conveyor-elevator was
built on the premises and consists of a 14-in. eight-ply rubber
belt, with buckets of sheet steel placed at intervals of 6 in.,
running over flanged pulleys. The buckets, or more properly
speaking the flights, are made from No. 12 steel plate, flanged to
produce the back and ends, with the ends secured to the flanged
bottom by one rivet in each. The plant has been in operation
for sixteen months and there have been few or no repairs to the
elevator, except to renew the belt, which is attacked by the acid
contained in the charges. This first belt was in continuous use
for nine months. As originally designed, the capacity was 100
tons per day of 12 hours, but this was found to require a speed
so high that the workmen were unable to handle the output of
the press. The speed was, consequently, reduced about 25 per
cent., which brings the output down to about 75 tons per day.
This output, as expressed in weight, naturally varies somewhat
owing to the variation in the weight of the material handled.

It is probable that the capacity could be increased to about
90 tons by enlarging the bricks, which could be done, but would
require a considerable amount of alteration in the machine, as
it is designed to produce a standard sized building brick. By
this method of increase, however, the work of handling would

SINTERING AND BRIQUETTING 67

not be materially increased, because the number of bricks would
be the same as with the present output of 75 tons; there would
be about 16 per cent, more to handle, by weight. Working on
the basis of 100 tons capacity, the bins were designed to afford
storage room for about three days' run, or a little over 300 tons.
The bins are made entirely of steel, in order that the hot material
may be dumped into them directly from the roasting furnaces,
thus saving one handling. In order that there may be room for
several kinds of material, the bins are divided into seven com-
partments, three on one side and four on the other. The lower
part is of f-in. steel plate, and the upper, about one-half the
hight, of ye -in. plate.

It may be well to call attention to the method of handling
the material, preparatory to its delivery to the brick press.
The bins are constructed, as will be seen by the drawing, with
their floor set 2.5 ft. above the working floor, which enables the
workmen to reach the material with a minimum effort. The
floor of the bins project 2.5 ft. in front of the face, thus forming
a platform on which the shoveling may be done without the
necessity of bending over. In this projecting platform are cut
rectangular holes 12 x 18 in., which are placed midway between
the openings in the front of the bins and furnished with screens
to stop any stray bolts or other coarse material that might injure
the press. This position of the holes through the platform was
adopted so that, in the event of the material running out beyond
the opening in the face, it would not fall directly upon the floor.
Two buckets are provided, with a capacity of 7 cu. ft. each,
which is the size of a single charge of the Chile mill. These buckets
have a hopper-shaped bottom fixed with a swinging gate which
is operated by the foot; thus the bucket can be run over the
pan of the Chile mill and the charge dumped directly into it.
The buckets run on an overhead iron track (1 in. by 3 in.) hung
7 ft. in the clear, above the floor.

The method of making up the charge is as follows : The bucket
is run under the hole in the platform nearest to the compartment
containing the material of which the charge is partly composed,
and a predetermined number of shovelfuls is drawn out and put
into the bucket, which is then pushed on to the next compartment
from which material is wanted, where the operation is repeated.
After charging into the bucket the requisite amount of ore or

LEAD SMELTING AND REFINING

FIG. 1 (a). — Plant for Bricking Ores, Selby Smelter. (Plan.)

flue dust, the bucket is run to the back of the building, where
the necessary amount of lime (slaked) is added. By putting the
lime in last, it is so surrounded by the dust or ore that it has
not the opportunity to stick to the sides of the bucket in discharg-
ing, as it otherwise would.

SINTERING AND BRIQUETTING

The number of men required to operate the entire plant,
exclusive of those employed in bringing the material to the bins
and emptying the cars into them, is 12, placed as follows: One
preparing the lime for use, one removing the charge from the
mill and supplying the elevator-conveyor, which is accomplished
by means of a specially shaped, long-handled shovel; one keeping
the supply spout of the press clear (an attempt was made to do
this mechanically, but was found to be unsuccessful, owing to
the extremely sticky nature of the material, and so was discarded
in favor of manual labor); one to control the press in case of

FIG. 1 (6). — Plant for Bricking Ores, Selby Smelter. (Elevation.)

mishap and to keep the dies clean; one oiler; three receiving the
bricks from the press and taking the brick-loaded cars from the
press to the drying-house, and two placing the bricks on the
shelves.

The drying-house scarcely requires description; it is but a
roofed shed, without sides, fitted with stalls into which the
bricks are set on portable shelves, as close as working conditions
will permit. The means of drying, at the present time, is by the
natural circulation of air, but a mechanical system is in contem-

70 LEAD SMELTING AND REFINING

plation, by which the air will be drawn into the building from
the outside and forced to find its way out through the bricks.
The drying-house is adjacent to the pressing plant, in fact forms
the back of it, so that there is a minimum distance to haul the
product. The time required for drying the bricks sufficiently
for them to withstand the necessary handling is, depending on
the weather, from two to eight days, the usual time being about
three days.

PART IV
SMELTING IN THE BLAST FURNACE

MODERN SILVER-LEAD SMELTING1

BY ARTHUR S. D WIGHT

(January 10, 1903)

The rectangular silver-lead blast furnace developed in the
Rocky Mountains has an area of 42 x 120 to 48 x 160 in. at
the tuyeres; 54 x 132 to 84 x 200 in. at the top; and hight from
tuyere level to top of charge of 15 to 21 ft. Such a furnace
smelts 80 to 200 tons of charge (ore and flux, but not slag and
coke) per 24 hours. The slag that has to be resmelted amounts
to 20 to 60 per cent, of the charge. Coke consumption is 12 to
16 per cent, of the charge. The blast pressure ranges from 1.5
to 4 Ib. per square inch, averaging close to 2 Ib. Gases of hand-
charged furnaces are taken off through an opening below the
charge-floor, the furnace being fed through a slot (about 20 in.
wide, extending nearly the whole length of the furnace) in the
iron floor-plates; or through a hood (brick or sheet iron) above
the charge-floor level, with a down-take to the flues, charge-doors
being provided on each side of the hood, extending preferably
the whole length of the furnace and usually having a sill a few
inches high which compels the feeder to lift his shovel.

When a silver-lead blast furnace is operating satisfactorily,
the following conditions should obtain: (1) A large proportion of
the lead in the charge should appear as direct bullion-product
at the lead- well. (2) The slag should be fluid and clean. (3) The
matte should be low in lead. (4) The furnace should be cool and
quiet on top, making a minimum quantity of lead-fume and
flue-dust, and the charges should descend uniformly over the
whole area of the shaft. (5) The furnace speed should be good.
(6) The furnace should be free from serious accretions and crusts;
that is to say, the tuyeres should be reasonably bright and open,

1 Abstract of portion of a paper presented at the Mexican meeting of
the American Institute of Mining Engineers, under the title "The Mechanical
Feeding of Silver-Lead Blast Furnaces." Transactions, Vol. XXXII, pp.
353-395.

74 LEAD SMELTING AND REFINING

and the level of the lead in the lead- well should respond promptly
to variations of pressure, caused by the blast and by the hight
of the column of molten slag and matte inside the furnace — an
indication that ample connection exists between the smelting
column and the crucible. Good reduction (using that term to
express the degree in which the furnace is manifesting its reducing
action) is obtained when the first three of the above conditions
are satisfied.

For any given furnace there are five prime factors, the resultant
of which determines the reduction, namely: (a) Chemical compo-
sition of the furnace charges; (6) proportion and character of
fuel; (c) air- volume and pressure, to which might perhaps also
be added temperature of blast; for, although hot blast has not
yet been successfully applied in lead-smelting practice, I believe
it is only a question of time when it will be; (d) dimensions and
proportions of smelting furnace; (e) mechanical character and
arrangement of the smelting column.

All but one of the above factors can be intelligently gaged.
The mechanical factor, however, can be expressed only in gener-
alities and indefinite terms. A wise selection of ores and proper
preliminary preparation, crushing the coarse and briquetting the
fine, will do much to regulate it, but all this care may be largely
nullified by careless feeding. The importance and possibilities
of the mechanical factor are generally overlooked and its symp-
toms are wrongly diagnosed. For instance, the importance of
slag-types has undoubtedly been considerably exaggerated at the
expense of the mechanical factor. Slags seldom come down
exactly as figured. We must know our ores and apply certain
empirical corrections to the iron, sulphur, etc., based on previous
experience with the ores; but these empirical corrections may
represent also an unformulated expression of the influence of the
mechanical factor on the reduction — a function, therefore, of the
ruling physical complexion of the ores, and the peculiarities of
the feeding habitually maintained in the works concerned. With
a given ore-charge large reciprocal variations may be produced
in the composition of slag and matte by merely changing the
mechanical conditions of the smelting column, and since the
efficient utilization of both fuel and blast must be controlled in
the same way, the mechanical factor may be considered, perhaps,
the dominating agent of reduction. Inasmuch as there is no

SMELTING IN THE BLAST FURNACE 75

way of gaging it, however, the only recourse is to seek a correct
adjustment and maintain it as a positive constant, after which
slag, fuel and blast may be with much greater certainty adjusted
toward efficiency of furnace work and metal-saving.

Behavior of Iron. — The output of lead is so dependent upon
the reactions of the iron in the charge that the chief attention
may well be fixed upon that metal as the key to the situation.
The success of the process depends largely upon reducing just
the right amount of iron to throw the lead out of the matte, the
remainder of the iron being reduced only to ferrous oxide and
entering the slag. Too much iron reduced will form a sow in
the hearth. Iron is reduced from its oxides principally by con-
tact with solid incandescent carbon, and by the action of hot
carbon monoxide. Reduction by solid carbon is the more waste-
ful, but there is in lead smelting an even more serious objection
to permitting the reduction to be accomplished by that means,
which leads to comparatively hot top and more or less volatiliza-
tion of lead. Reduction by carbon monoxide is the ideal condi-
tion for the lead furnace. It means keeping the zone of incan-
descence low in the charge column, leaving plenty of room above
for the gases to yield up their heat to, and exercise their reducing
power on, the descending charge, so that by the time they escape
they will be well-nigh spent. Their volume and temperature will
be diminished, and the low velocity of their exit will tend to
minimize the loss of lead in fume and flue dust.

The idea that high temperatures in lead blast furnaces should
be avoided is based on a misconception. Temperatures must
exist which are sufficiently high to volatilize all the lead in the
charge, if other conditions permit. A high temperature before
the tuyeres means fast smelting; and fast smelting, under proper
conditions, means a shortening of the time during which the lead
is subject to scorifying and volatilizing influences. A rapidly
descending charge, constantly replenished with cold ore from
above, absorbs effectively the heat of the gases and acts as a
most efficient dust and fume collector. In considering long flues,
bag-houses, etc., it should be kept in mind that the most effective
dust collector ought to be the furnace itself.

In the practice of twelve years ago and earlier, particularly
when using mixed coke and charcoal, reduction by carbon was
probably the rule; and the percentage of fuel required was very

76 LEAD SMELTING AND REFINING

high. There is good reason to think we have still much room
for improvement along this line in our average practice of today.

Volume of Blast. — It is customary to supply a battery of
furnaces from a large blast main, connected with a number of
blowers. Inasmuch as the air will take preferably the line
of least resistance, if the internal resistance of any one furnace
be increased the volume of air it will take will be diminished
and the others will be favored unduly. Only by keeping all the
furnaces on approximately the same charge, with the same hight
of smelting column, can anything like uniformity of operation
and close regulation be secured. The rational plan would seem
to be to have a separate blower, of variable speed, directly con-
nected to each furnace, but this plan, which has had a number
of trials, has usually been abandoned in favor of the common
blast main. Trials by myself, extending over considerable
periods, have been so uniformly favorable, however, that I am
forced to ascribe the failure of others to some outside reason.

The peculiar atmosphere required in the lead blast furnace
depends upon the correct proportion of two counteractive ele-
ments, carbon and oxygen. If given too much air the furnace
will show signs of deficient reduction, commonly interpreted as
calling for more fuel, which will be sheer waste since its object
is to burn up surplus air. There will be an additional waste
through the extra coal burned under the steam boilers. The
true remedy would be to cut down the quantity of air. Burning
up excessive coke is as hard work as smelting ore. Too much
fuel invariably slows up a furnace; it also drives the fire upward
and gives predominance to reduction by solid carbon. The
maintenance of a minimum fuel percentage, with a correctly
adjusted volume of air, will tend to promote the conditions under
which iron will be reduced by the gases, rather than by solid
carbon.

Pressure of Blast. — Pressure necessarily involves resistance;
and the blast-pressure, as registered by a simple mercury-gage
on the bustle-pipe, may be increased in two ways: (1) By increasing
the volume of air forced through the interstices in the charge.
This is the wrong way; but, unfortunately, it is only too common in
our practice, and therefore deserves to be mentioned, if only to
be condemned. (2) By leaving the volume of air unchanged, but
increasing the friction offered by the interstitial channels, either

SMELTING IN THE BLAST FURNACE 77

by making them smaller in aggregate cross-section (which means
a finer charge), or by making them longer (which means a higher
smelting column). A correctly graduated internal resistance is,
therefore, the only true basis for a high blast furnace, which,
when so produced, will bring about rapid smelting, a low zone of
incandescence, and a very vigorous action upon the ores by the
gases in their retarded ascent through the charge column. These
conditions promote the reduction of iron by CO. The adjustment
of internal resistance, which is thus clearly the main factor, can
be accomplished only by the correct feeding of the furnace.

Feeding the Charge. — It is self-evident that, the more thor-
ough the preliminary preparation of the charge before it reaches
the zone of fusion, the more rapidly can the actual smelting
proceed. A piece of raw ore that finds itself prematurely at the
tuyeres, without having been subjected to the usual preparatory
processes of drying, heating, reduction, etc., must remain there
until it is gradually dissolved or carried away mechanically in
the slag. Any such occurrence must greatly retard the process.
It would seem, by the same reasoning, that an intimate mixture
of the ingredients of the charge should expedite the smelting,
and I advocate the intimate mixture of the charge ingredients
in all cases.

The theory of feeding is simple, but not so the practice. If
the charge column were composed of pieces of uniform size, the
ascending gases would find the channel of least resistance close
to the furnace walls and would take it preferably to the center
of the shaft. The more restricted channel would necessitate a
higher velocity, so that not only would the center of the charge
be deprived of the action of the gases, but also the portion trav-
ersed would be overheated; many particles of ore would be
sintered to the walls or carried off as flue dust; slag would form
prematurely; fuel would be wasted; in short, all the irregularities
and losses which accompany over-fire would be experienced. In
practice the charge is never uniform, but is a mixture of coarse
and fine. By lodging the finer material close to the walls and
placing the coarser in the center, an adjustment may be made
which will cause the gases to ascend uniformly through the
smelting column. A furnace top smoking quietly and uniformly
over its whole area is the visible sign of a properly fed furnace.

Effect of Large Charges. — It has frequently been remarked

78 LEAD SMELTING AND REFINING

that, within certain limits, large charges give more favorable
results than small ones; and numerous attempts have been made
to account for this fact. My observations lead me to offer the
following as a rational explanation — at least in cases where ore
and fuel are charged in alternate layers. Large ore-charges mean
correspondingly large fuel-charges. The gases can pass readily
through the coke; and hence each fuel-zone tends to equalize
the gas currents by giving them another opportunity to distribute
themselves over the whole furnace area, while each layer of ore
subsequently encountered will blanket the gases, and compel
them to force a passage under pressure, which is the manner
most favorable to effective chemical action.

In mechanically fed furnaces the charges of ore and fuel are
usually dropped in simultaneously from a car and the separate
layers thus obliterated, and the distributing zones which are
•such a safeguard against the consequences of bad feeding are
lacking, hence more care must be exercised to secure proper
placing of the coarse and fine material. This may throw some
light on the failure of most of the early attempts at mechanical
feeding.

Mechanical Character of Charge. — Very fine charges blanket
the gases excessively and cause them to break through at a few
points, forming blow-holes, which seriously disturb the operation,
cause loss of raw ore in the slag, and are accompanied by all the
evils of over-fire. A charge containing a few massive pieces,
the rest being fine, is a still more unfavorable combination. A
very coarse charge permits too ready an exit to the gases, and in
the end tends likewise to over-fire and poor reduction. The
remedy is to briquette the fine ore (though preferably not all of
it), and crush the coarse to such degree as to approach an ideal
result, which may be roughly described as a mixture in which
about one-third is composed of pieces of 5 to 2 in. in diameter,
one-third pieces of 2 to 0.5 in., and the remaining third from
0.5 in. down. The coke is better for being somewhat broken up
before charging, and a reasonable amount of coke fines, such as
usually accompanies a good quality of coke, is not in the least
detrimental. The common practice of handling the coke by
forks and throwing away the fines is to be condemned as an
unwarranted waste of good fuel. The slag on the charge should
be broken to pieces at most 6 in. in diameter. The common

SMELTING IN THE BLAST FURNACE 79

practice of throwing in whole butts of slag-shells is bad. There
is no economy in using the slag hot; cold charges, not hot, are
what we want. A reasonable amount of moisture in the charge
is beneficial, providing it be in such form as to be readily dried
out. It is often advantageous to wet the ore mixtures while
bedding them, or to sprinkle the charges before feeding. The
driving off of this water must consume fuel, but not so much as
if the smelting zone crept up. Large doses of water applied
directly to the furnace are unpardonable under any circumstances,
however, though they are sometimes indulged in as a drastic
measure to subdue excessive over-fire when other and surer
means are not recognized. One of the chief merits of moderate
sprinkling before charging is that it gives in many cases a more
favorable mechanical character, approximating a lumpy condition
in too fine a charge, and assisting to pack a too coarse one.

Different Behavior of Coarse and Fine Ore. — In taking up a
shovelful of ore, the fine will be observed to predominate in the
bottom and center, and the coarse on the top and sides. When
thrown from the shovel, the coarse will outstrip the fine and fall
beyond it. In making a conical pile the coarse ore will roll to
the base, leaving the fine near the apex. This difference in the
action of the mobile coarse ore and the sluggish fines is the key
to the practical side of feeding, both manual and mechanical.
It is not sufficient to tell the feeder to throw the coarse in the
middle and the fine against the sides; if it be easier to do it some
other way such instructions will count for little. The desired
result can be best secured by making the right way easier than
the wrong way.

It is generally conceded that the open-top furnaces, fed by
hand through a slot in the floor-plates, do not give as satisfactory
results as the hooded furnaces with long feed-doors on both
sides. In the open-top furnace it is comparatively difficult to
throw to the sides; the narrower the slot the greater the difficulty.
The major part of the charge will drop near the center, making
that place higher than the sides. The fine ore will tend to stay
where it falls, while the coarse will tend to roll to the sides, thus
leading to an arrangement of the charge just the reverse of what
it ought to be. In the hooded furnace most of the material will
naturally fall near the doors, causing the sides to be higher than
the center toward which the coarse will roll, while the force of

80 LEAD SMELTING AND REFINING

the throw as the ore is shoveled in will also have a tendency to
concentrate the coarse material in the center.

Once a proper balance of conditions has been found, absolute
regularity of routine is the secret of good results. An experienced
and intelligent feeder owes his merit to his conscientious regularity
of work. He may have to vary his program somewhat when
he encounters a furnace that is suffering from the results of bad
feeding by a predecessor; but his guiding principle is first to
restore regularity, and then maintain it. A poor feeder can
bring about, in a single shift, disorders that will require many
days to correct, if indeed they are corrected at all during the
campaign. The personal element is productive of more harm
than good.

Mechanical Feeding. — If it be admitted that the work of a
feeder is the better the more it approximates the regularity of
that of a machine, it ought to be desirable to eliminate the per-
sonal factor entirely and design a machine for the purpose, which
would be a comparatively simple matter if it be known just
what we want to accomplish. No valid ground now exists for
prejudice against mechanical feeding in lead smelting. It is in
successful operation in a number of large works, and is being
installed in others. Our furnaces have outgrown the shovel; we
have passed the limit of efficiency of the old methods of handling
material for them. We must come to mechanical feeding in
spite of ourselves. But whatever may be the motive leading to
its introduction, its chief justification will be discovered, after it
has been successfully installed and correctly adjusted, in the
consequent great improvement of general operating results, metal
saving, etc. It will remove one of the most uncertain factors
with which the metallurgist has to deal, thereby bringing into
clearer view for study and regulation the other factors (fuel and
blast proportion, slag composition, etc.) in a way that has hardly
been possible under the irregularities consequent upon hand
feeding.

MECHANICAL FEEDING OF SILVER-LEAD BLAST
FURNACES *

BY ARTHUR S. DWIGHT

(January 17, 1903)

Historical. — A silver-lead furnace fed by means of cup and
cone was in operation in 1888 at the works of the St. Louis Smelt-
ing and Refining Company at St. Louis, Mo., but it is probable
that previous attempts had been made, since Hahn refers (" Min-
eral Resources of the United States," 1883) in a general way to
experiments with this device, which were unsuccessful because
the heat crept up in the furnace and gave over-fire. At the time
of my visit to the St. Louis works (in 1888) the furnaces were
showing signs of over-fire, but this may not have been their
characteristic condition. A. F. Schneider, who built the St.
Louis furnaces, afterward erected, at the Guggenheim works at
Perth Amboy, N. J., round furnaces with cup and cone feeders,
but although good results are said to have been obtained, the
running of refinery products is no criterion of what they would
do on general ore smelting,

Cup and Cone Feeders. — The cup and cone is an entirely
rational device for feeding a round furnace, but is quite unsuitable
for feeding a rectangular one. Furnaces of the latter type were
installed for copper smelting at Aguas Calientes, Mex., with two
sets of circular cup and cone feeders, but disastrous results fol-
lowed the application of this device to lead furnaces. The reason
is clear when it is considered that a circular distribution cannot
possibly conform to the requirements of a rectangular furnace.
A more rational device was designed for the works at Perth
Amboy, N. J.

Pfort Curtain. — About ten years ago some of the American
smelters adopted the Pfort curtain, which, as adapted to their

1 Abstract of a paper ("The Mechanical Feeding of Silver-Lead Blast
Furnaces") presented at the Mexican meeting of the American Institute of
Mining Engineers and published in the Transactions, Vol. XXXII. For the
first portion of this paper see the preceding article.

LEAD SMELTING AND REFINING

requirements, consisted of a thimble of sheet iron hung from the
iron deck plates so as to leave about 15 in. of space between it
and the furnace walls, this space being connected with the down-
take of the furnace. The thimble was kept full of ore up to the

FIG. 2. — Perth Amboy, N. J., Lead Fur-
nace. Vertical section at right angles to
Fig. 3.

charge-floor. This device was popular for a time, chiefly because
it prevented the furnace from smoking and diminished the labor
of feeding, but it was found to give bad results in the furnaces,
it being impossible to observe how the charge sunk (except by

SMELTING IN THE BLAST FURNACE

dropping it below the thimble), while the curtain had to be
removed in order to bar down accretions, and, most important,
it caused irregular furnace work and high metal losses, because it

FIG. 3. — Perth Amboy, N. J., Lead Furnace. Vertical section at right

angles to Fig. 2.

effected a distribution of the coarse and fine material which was
the reverse of correct, the evil being emphasized by the taking
off of the gases close to the furnace walls.

84 LEAD SMELTING AND REFINING

Terhune Gratings. — R. H. Terhune designed a device (United
States patent No. 585,297, June 29, 1897), which comprised two
grizzlies, one on each side of the furnace, sloping downward from
the edge of the charge-floor toward the center line of the furnace.
The bars tapered toward the center of the furnace, the open
spaces tapering correspondingly toward the sides, so that as the
charge was dumped on them a classification of coarse and fine
would be effected. This device is correct in conception.

Pueblo System. — In the remodeling of the plant of the Pueblo
Smelting and Refining Company in 1895, under the direction of
W. W. Allen, mechanical feeding was introduced, and the system
was the first one to be applied successfully on a large scale. The
furnaces of this plant are 60 x 120 in. at the tuyeres, with six
tuyeres, 4 in. in diameter on each side, the nozzles (water cooled)
projecting 6 in. inside the jackets. The hight of the smelting
column above the tuyeres is 20 ft. The gases are taken off
below the charge-floor, and the furnace tops are closed by hinged
and counter-weighted doors of heavy sheet iron, opened by the
attendant, just previous to dumping the charge-car. In the side
walls of the shaft are iron door-frames, ordinarily bricked up,
but giving access to the shaft for repairs or barring out without
interfering with the movement of the charge-car. Extending
across the shaft, about 18 in. above the normal stock line, are
three A-shaped cast-iron deflectors, dividing the area of the
shaft into four equal rectangles.

The general arrangement of the plant is shown in Fig. 4.
From the charge-car pit there extends an inclined trestle, on an
angle of 17 deg. to the charge-floor level, in line with the battery
of furnaces. The gage of the track is approximately equal to
the length of the furnaces at the top. The charge-car, actuated
by a steel tail-rope, moves sideways on this track from the charg-
ing-pit to any furnace in the battery. The hoisting drums are
located at the crest of the incline, inside of the furnace building.
At the far end of the latter there is a tightener sheave, with a
weight to keep proper tension on the tail-rope. The charge-car
has a capacity of 5 tons. It has an A-shape bottom, and is so
arranged that one attendant can quickly trip the bolt and discharge
-the car.

While the car is making its trip the charge-wheelers are filling
their buggies, working in pairs, each man weighing up a half-

LEAD SMELTING AND REFINING

charge of a particular ingredient. They then separate, each
taking his proper place in the line of wheelers on either side.
When the car has returned, the wheelers successively discharge
their buggies into opposite ends of the car. The coke is added
last, to avoid crushing. The system is not strictly economical
of labor, since the wheelers, who must always be ready for their
car, have to wait for its return, which necessitates more wheelers
than would otherwise be required. Figs. 5, 6 and 7 show the car.
A vertical section through the car filled by dumping from the
two ends will show an arrangement of coarse and fine, which is
far from regular. Analyzing its structure, we shall find a conical
pile near each end, with a valley between them, in which coarse

FIG. 5. — Pueblo Charge-car. (Side elevation.)

ore will predominate. The deflectors in the furnace, previously
referred to, serve to scatter the fines as the charge is dropped in.
Without them the feeding of the furnace would be a failure; with
them it is successful, though not so completely as might be, the
furnaces having a tendency to run with hot tops. With the
battery of seven furnaces, each smelting an average of 100 tons
of ore per day, the saving, as compared with hand-feeding, was
$63 per day, or 9c. per ton of ore, this including cost of steam,
but not wear and tear on the machinery. This is distinctly a
maximum figure; with fewer furnaces the fixed charges of the
mechanical feed would soon increase the cost per ton to such a
figure that the two systems would be about equal in economy.
East Helena System. — This was introduced at the East

SMELTING IN THE BLAST FURNACE

R n

-J=i-L_ _____.w_ JJrrj.

FIG. 6. — Pueblo Charge-car. (Plan.)

* //

0
0

„/

^J, ,

a
o]

^ _ E

r — - -^

[^u_^ <fi« — -fly/^ N>K>

— * SSS

-10-0-

FIG. 7. — Pueblo Charge-car. (End elevation.)

88 LEAD SMELTING AND REFINING

Helena plant of the United Smelting and Refining Company by
H. W. Hixon. The plant comprised four lead furnaces, each
48 x 136 in., with a 21-ft. smelting column. They were all open-
top furnaces, fed through a slot over the center, the gases being
taken off below the floor. They were capable of smelting about
180 tons of charge (ore and flux) per 24 hours, using a blast of
30 to 48 oz., furnished by two Allis duplex, horizontal, piston
blowers, air-cylinders 36 in. diam., 42-in. stroke, belted from
electric motors. The Hixon feed was designed to meet existing
conditions, without irrevocably cutting off convenient return to
hand feeding in case of an emergency. As shown in Fig. 9 there
is a track-way at right angles to the line of furnaces. The car
hoisted up the incline is landed on a transfer carriage, on which,
-after detaching the cable, it can be moved over the tops of the

FIG. 8. — Pueblo System. (Sectional diagrams of furnace top.).

furnaces by means of a tail-rope system. The gage of the charge-
car is 4 ft. 9 in.; of the transfer carriage, 11 ft. 8 in. A switch at
the lower end of the incline permits two charge-cars to be em-
ployed, one being filled while the other is making the trip. In
sending down the empty car a hand winch is necessary to start
it from the transfer carriage. Figs. 10 and 1 1 show the charge-car;
Fig. 12 the transfer carriage.

The charge-car is 10 x 4 x 3.5 ft., and has capacity for 6 tons
of ore, flux, slag and fuel, the total of ore and flux being usually
8800 Ib. Its bottom is flat, consisting of two doors, hinged along
the sides and kept closed by means of chains wound about a
longitudinal windlass on top of the car. The charging pits
-are decked with iron plates, leaving a slot along the center of

SMELTING IN THE BLAST FURNACE

each car exactly like the slot in the furnace top. The loaded
ore-buggies are taken from the wheelers by two men, who care-
fully distribute the contents of each buggy along the whole length

FIG. 9. — East Helena System. (Vert-longitudinal section and plan of incline.)

of the charge-car by dragging it along the slot while in the act
of dumping. Each buggy contains but one ingredient; they
follow one another in a prescribed order, so as to secure thin
layers in the charge-car. The coke is divided into three or
more layers.

FIG. 10. — East Helena Charge-car. (Side elevation.)

The first few trials of this device were not satisfactory. The
furnaces quickly showed over-fire, and decreased Iea4 out put ,
which would not yield to any remedy except a return to hand

LEAD SMELTING AND REFINING

feeding. The total charge being dropped in the center of the
furnace, a central core of fines was produced, the lumps tending
to roll toward the walls. This wrong tendency was emphasized
by the presence of the chains supporting the bottom of the charge-
car. On unwinding them to dump the car, the doors were pre-
vented from dropping by the wedging of the chains in the charge,
which in turn arched itself more or less against the sides of the
car; hence the doors opened but slowly, and often had to be
assisted by an attendant with a bar. In consequence of this
slow opening, considerable fine ore sifted out first and formed a
ridge in the center of the furnace, from the slopes of which the
coarser part of the charge, the last to fall, naturally rolled toward
the sides. This fact, determined during a visit of the writer in

FIG. 11. — East Helena Charge-car. (Plan.)

April, 1899, proved to be the key to the situation. The attendant
operating the tail-rope mechanism was instructed to move the
transfer carriage rapidly backward and forward over the slot
while the first one-third or one-half of the charge was dropping,
and during the rest of the discharge to let the car stand directly
over the slot and permit the coarser material to fall in the center
of the furnace. Two piles of comparatively fine material were
thus left on the charge-floor, one on each side of the slot. These
were subsequently fed in by hand, with instructions to throw the
material well to the sides of the furnace.

The furnaces were running very hot on top when this modified
procedure was begun. In a few hours the over-fire had disap-
peared; the lead output was increasing; and the furnaces were
running normally. This was done about May 1, 1899, and from

SMELTING IN THE BLAST FURNACE

that time until about February 20, 1900, the Hixon feed, as
modified above, was continuously in operation. In October, 1898,
with three furnaces in operation and hand feeding, the labor cost
per furnace was $42.06 per day; in October, 1899, with the same
number of furnaces and mechanical feeding, it was $41 per day,
the saving being only 0.6c. per ton of charge.

Dwight Spreader and Curtain. — In January, 1900, the writer
again had occasion to visit the East Helena plant, to investigate
why a certain cheap local coke could not be used successfully
instead of expensive Eastern coke. Strange as it may seem, the

FIG. 12. — East Helena Charge-car and Transfer Carriage. (Elevation.)

peculiar behavior of the cokes was traced to improper feeding
of the furnaces. Further study of the mechanical feeding system,
then in operation for nine months, showed that it was far from per-
fect, and it appeared desirable to design a spreader which would
properly distribute the material discharged from the Hixon car
and dispense with hand feeding entirely. An experimental con-
struction was arranged, as shown in Fig. 13. The flanged cast-
iron plates around the feeding slot were pushed back and a
roof-shaped spreader, with slopes of 45 deg., was set in the gap,
leaving openings about 8 in. wide on each side. The plan pro~

LEAD SMELTING AND REFINING

vided for two iron curtains to be hung, one on each side of
the spreader, and so adjusted that the fine ore sliding down the
spreader would clear the edge of the curtain and shoot toward
the sides of the furnace, while the coarse ore would strike the cur-
tain and rebound toward the center of the furnace. The classi-
fication effected in this manner was capable of adjustment by
raising or lowering the curtain. This arrangement was found to

FIG. 13. — East Helena System, with spreader and cur-
tains. (Experimental form.)

work surprisingly well. The first furnace equipped with it imme-
diately showed improvement. It averaged better in speed, with
lower blast, lower lead in slag and matte, and better bullion
output than the other furnaces operating under the old system.
The success of the spreader and curtain being established, the
furnaces were provided with permanent constructions, the only
modifications being that the ridge of the spreader was lowered
to correspond with the level of the floor and the curtains were

SMELTING IN THE BLAST FURNACE

omitted, the feeding being apparently satisfactory without their
aid. In their absence, the lowering of the spreader was a proper
step, as it distributed the material fully as well, and caused less
abrasion of the walls. The final form is shown approximately in
Fig. 14. It has given complete satisfaction at East Helena since
February, 1900, and has been adopted as the basis for the mechan-
ical feeding device in the new plant of the American Smelting
and Refining Company at Salt Lake, Utah.

Comparison of Systems. — In mechanical design the Pueblo

FIG. 14. — East Helena System.
approximate.)

(Final form,

system is better than the East Helena, being simpler in construc-
tion and operation. No time is lost in attaching and changing
cables, operating transfer carriage, etc. In both systems the
track runs directly over the tops of the furnaces, and this is an
inconvenience when furnace repairs are under way. The Pueblo
car is the simpler, and makes the round trip in about half the
time of a car at East Helena, so the two cars of the latter do not
make much difference in this respect. The system of filling the
charge-car at Pueblo is also the quicker. It may be estimated

94 LEAD SMELTING AND REFINING

roughly that per ton of capacity it takes 2.5 to 3 times as long
to fill the East Helena car; and this means longer waiting on the
part of the wheelers, and consequently greater cost of moving
the material, representing probably 7 or 8c., in favor of Pueblo,
per ton of charge handled. However, both systems are wasteful
of labor. As to furnace results, it is believed that the better
distribution of the charge in the East Helena system leads to
greatly increased regularity of furnace running, less tendency to
over-fire, some economy in fuel, less accretions on the furnace
walls and larger metal savings. If the half of these conclusions
are true, the difference of 7 or 8c. per ton in favor of the Pueblo
system, which can be traced almost entirely to the cost of filling
the charge-car, sinks into insignificance in comparison with the
important advantages of having the furnaces uniformly and
correctly fed.

True Function of the Charge-Car. — The radically essential
feature of a mechanical feeding device is that part which auto-
matically distributes the material in the furnace, whatever
approximate means may have been used to effect the delivery.

Taking a hasty review of the numerous feeding devices that
have been tried in lead-smelting practice, we cannot but remark
the fact that those which depended upon dumping the charge
into the furnace from small buggies or barrows failed generally
to secure a proper classification and distribution of coarse and
fine, and, consequently, were abandoned as unsuccessful, while
the adoption of the idea of the charge-car for transporting the
material to the furnace in large units seems to have been coinci-
dent with a successful outcome. It is natural enough, therefore,
that the car should be regarded by many as the vital feature.
This view of the question is not, however, in accordance with the
true perspective of the facts, and merely limits the field of appli-
cation in an entirely unnecessary way. It must be apparent
that the essential function of the charge-car is cheap and con-
venient transportation. The distribution of the charge is an
entirely different matter, in which, however, the charge-car may
be made to assist, as in the Pueblo system; or entirely distinct
and special means may be employed for the distribution, as in
the East Helena system.

To follow the argument to its conclusion, let us imagine for
the moment that the East Helena plant were arranged on the

SMELTING IN THE BLAST FURNACE 95

terrace system, with the furnace tops on a level with the floor
of the ore-bins. Certain precautions being observed, the spreader
would give as good results with small units of charge delivered
by buggies as it now does with the large units delivered by the
charge-car, and the expense of delivery to the furnaces would be
practically no more than it now is to the charge-car pit. The
furnace top would, of course, have to be arranged so that the
buggies, in discharging, could be drawn along the slot, so as to
give the necessary longitudinal distribution parallel to the furnace
walls, just as is now done in filling the charge-car. The ends of
the spreader, if built like a hipped roof, would secure proper
feeding of the front and back.

Thus, by eliminating the charge-car, and with it the necessity
for powerful hoisting machinery, with its expensive repairs and
operating costs, we may greatly simplify the problem of mechan-
ical feeding, and open the way for the adoption of successful
automatic feeding in many existing plants where it is now con-
sidered impracticable.

COST OF SMELTING AND REFINING

BY MALVEBN W. ILES

(August 18, 1900)

In the technical literature of lead smelting there is a lament-
able lack of data on the subject of costs. The majority of writers
consider that they have fulfilled their duties if they discuss in
full detail the chemical and engineering sides of the subject,
leaving the industrial consideration of cost to be wrought out by
experience. When an engineer or metallurgist collects data on
the costs involved in the various smelting operations, he generally
hesitates to give this special information to the public, as he
regards it as private, or reserves it as stock in trade to be held
for his own use.

The following tables of cost have been compiled from actual
results of smelting and refining at the Globe works, Denver,
Colo., and are offered in the hope that they will prove a valuable
addition to the literature of lead smelting. These results are
offered tentatively, and, while true for the periods stated, they
require considerable adjustment to meet the smelting conditions
of the present time.

COST OF HAND-ROASTING PER TON (2000 LB.) OF ORE

1887 $3.975

1888 4.280

1889 4.120

1890 . . . 3.531

1891 $3.530

1892

1893

1894 . , 3.429

1895 $2.806

1896 2.840

1897 2.740

1898 . . 2.620

At first the roasting was done mainly by hand roasters; later
two Brown-O'Harra mechanical furnaces were used, and the
cost was reduced, but not to the extent usually conceded to this
type of furnace, as the large amount of repairs and the conse-
quent loss of time diminished the apparent gain due to greater
output. The figures quoted above may be considered somewhat
higher than the average, as the roasters were charged in propor-
tion with expenses of general management, office, etc.

SMELTING IN THE BLAST FURNACE 97

In viewing the yearly reduction of costs one must take into
consideration many changes in the furnace construction and
working, as well as the items of labor, fuel, etc. From 1887 to
1899 the principal changes in the construction of the hand-roasting
furnaces consisted in an increase of width, 2 ft., which allowed
an addition of 200 Ib. to each ore charge, and corresponded to a
total increase per furnace of 1200 Ib. in 24 hours. In the working
of the charge an important change was made in the condition of
the product. Formerly the material was fused in the fusion-box
and drawn from the furnace in a fused or slagged condition; and
while this gave an excellent material for the subsequent treatment
in the shaft furnace in that there was very little dusting of the
charge, and a considerable increase in the output of the furnace,
the disadvantages of large losses of lead and silver greatly over-
balanced the advantages, and called for an entire abandonment
of the fusion-box. As a result of experience it was found that
the best condition of product is a semi-fused or sintered state,
in which the particles of roasted ore have been compressed by
pounding the material, which has been drawn into the slag pots,
with a heavy iron disk. The amount of "fines" under these
conditions is quite small and depends upon the percentage of
lead in the ore, the degree of heat employed, and the extent of
the compression.

The total cost was partly reduced from the lessened labor
cost following the financial disturbance of 1893, and partly from
the reduction in the fuel cost, the former expensive lump coal
being replaced by the slack coals from southern Colorado.

The comparison of the cost of labor by the two methods
shows a gain of 54c. a ton in favor of the mechanical furnaces.
However, I consider that this gain is a costly one, and is more
than offset by the large amount of high-grade fuel required, and
the expense of repairs not shown in the following table. Indeed,
I believe that at the end of five or ten years the average cost
of roasting per ton by the hand roasters will be even smaller than
by these mechanical roasters.

To illustrate the details of roasting cost and to furnish a com-
parison of the hand roasters and mechanical furnaces, the following
table has been prepared:

LEAD SMELTING AND REFINING

DETAILS OP AVERAGE MONTHLY COST FOR 1898 OF HAND
ROASTERS AND MECHANICAL FURNACES

MONTH

HA>

D ROAST

ERS «
X cfl

• BRO
MECHA

wx-O'HA

«CAL Fu

RRA •
RNACES

TOTAL TONS
ROASTED

TONS
ROASTED
PER DAY

5,691
5,677
5,821
5,472
5,444
4,859
5,691
5,910
5,677
6,254
6,291
5,874

184
203
188
182
176
162
184
191
189
202
213
198

$1.47
1.44
.51
.47
.55
.58
.59
.55
.55
.48
.42
.45

$0.53
0.44
0.53
0.47
0.51
0.48
0.48
0.46
0.45
0.49
0.47
0.48

$0.80
0.99
0.64
0.71
0.84
0.71
0.75
0.83
0.74
0.72
0.80
0.78

$0.92
0.72
0.76
0.80
0.80
0.90
0.72
0.72
0.73
0.65
0.66
0.79

$0.80
0.58
0.64
0.69
0.69
0.68
0.56
0.55
0.55
0.50
0.53
0.63

$1.32
1.01
0.62
0.87
0.81
1.17
0.64
0.75
0.67
0.60
0.70
0.81

April

May

July

October

December

$1.50

$0.48

$0.77
2.75

$0.76

$0.62

$0.83
2.21

Total

Cost of Smelting. — The lead-ore mixtures of the United
States, in addition to lead, contain gold, silver and generally
copper, and are treated to save these metals. The total cost of
smelting is made up of a large number of items. The questions
of locality and transportation, fuel, fluxes and labor are the
principal factors, to which must be added the handling of the
material to and from the furnace; the furnace itself, its size,
shape, and method of smelting, the volume and pressure of blast,
etc. The following table of costs, from 1887 to 1898, shows in
a general way the great advance that has been made in the
development of smelting, and the consequent reduction in cost
per ton of ore treated:

AVERAGE COST OF SMELTING, PER TON

1887 $4.644

1888 4.530

1899 4.480

1890 . , ... 4.374

1891 $4.170

1892 4.906

1893 . . 3.375

1894 3.029

1895 $2.786

1896
1897
1898

2.750
2.520
2.260

In connection with this table of smelting cost should be con-
sidered the changes developed during the interval 1887-1889,
outlined as follows:

SMELTING IN THE BLAST FURNACE

CONDITIONS OF SMELTING IN 1886 AND 1899 CONTRASTED TO
SHOW THE PROGRESS OF DEVELOPMENT

gs .

* s£

i OS

^§

. |

£H*

°*rn:

« ^ ^

<: -

* Pk

s<s

IIs

H W W

W o «
2«g

*||

«1

»g£

w <

C/3

1886

30X100

In pots

Charcoal

By hand

280

200

By locomo-

1 By horse

1899.. . .

42 X 140

3to4

128

In furnaces

Coke

tive

3000-6000

j 2000-3000

I believe that there is room for further improvement in the
substitution of mechanical transportation within the works for
hand labor, and that the fuel cost can be materially reduced by
replacing the coke, which at present contains 16 to 22 per cent,
of ash, by a fuel of purer and better quality.

Cost of Refining by the Parkes Process. — In general it may
be stated that the average cost of refining base bullion is from
$3 to $5 a ton. This amount is based on the cost of labor, spelter,
coal, coke, supplies, repairs and general expenses. When the
additional items of interest, expressage, brokerage and treatment
of by-products are considered, which go to make up the total
refining cost, the amount may be stated approximately as $10
per ton of bullion treated.

Variations in the cost occur from time to time, and are due to
several causes, principally the irregularity of the bullion supply
and its consequent effect on the work of the plant. When the
amount of bullion available for treatment is small, the plant
cannot be run to its maximum capacity, and the cost per ton
will naturally be increased. To illustrate this variation, the
average cost per ton of base bullion refined during nine months
in 1893 was:

January, $4.864; February, $5.789; March, $5.024; April,
$3.915; May, $5.094; June, $4.168; July, $4.231; August, $4.216;
September, $5.299.

The yearly variation shows but little change, as the average
cost per ton was for 1893, $4.75; for 1894, $3.99; for 1895, $4.21;
for 1896, $3.90. In considering the total cost of refining, the
additional factors of interest, expressage, parting, brokerage, and
reworking of by-products must be considered. As the dor£ silver
is treated at the works or elsewhere, so will the total cost be less

100

LEAD SMELTING AND REFINING

or greater. The following table gives the cost in detail, when
the parting is done at the same works:

AVERAGE MONTHLY COST OF REFINING PER TON OF BULLION

TREATED

ITEMS

1895
JAN. TO JULY

1895
JULY TO DEC.

1896
JAN. TO JULY

AVERAGE

Labor

$2.351

$1.718

$1.836

$1.968

Spelter

0.757

0.840

0.987

0.861

Coal

0.585

0.442

0.461

0 496

Coke

0.634

0.418

0.511

0.521

Supplies, repairs and
general expenses ....
Interest

0.343

1.808

0.273
1.075

0.252
1.070

0.289
1.317

Expressage
Parting and brokerage .
Reworking by-products

1.360
2.483
1.567

1.015
2.084
1.286

0.882
1.796
1.625

1.085
2.121
1.492

Totals

$11.888

$9.151

$9.420

$10.151

Tons bullion refined. . . .

5,511.58

9,249.07

10,103.43

8,287.99

An analysis of the different items of cost is important, and a
brief summary is given below.

Labor and Attendance. — The cost for this item varies but little
from year to year, and its reduction depends, for the most part,
on a larger yield per man rather than on a reduction of wages.
If a man at the same or slightly increased cost can give a larger
output, so will the labor cost per ton be diminished. This result is
accomplished by enlarging the furnace capacity and by using ap-
pliances which will handle the bullion and its products in an easier
and quicker manner. The small size of the furnaces, settlers and
retorts used at modern refineries is open to criticism; I believe
that great improvement can be made in this direction.

Spelter. — The cost of this item varies with the market con-
ditions, and will probably be changed but little in the future,
as the amount necessary per ton of bullion seems to be fixed.

Coal. — The amount required per ton of bullion is fairly
constant, and while lessened cost for fuel may be attained by the
substitution of oil or gaseous fuel, the fuel cost in comparison
with the aggregate cost is very small, and leaves little opportunity
for improvement in this line.

Supplies. — This item includes brooms, shovels, wheelbarrows,
etc., and the amount is small and fairly constant from year to year.

Repairs. — This item is quite small in works properly con-

SMELTING IN THE BLAST FURNACE 101

structed; and in this connection I wish to call particular attention
to the floor covering, which should be made of cast-iron plates
from 1.5 to 2 in. thick, and placed on a 2- to 3-in. layer of sand
spread over the well-tamped and leveled ground. The constant
patching of brick floors is not only an annoyance, but is costly
from the additional labor required. Furthermore, a brick floor
does not permit a close saving of the metallic scrap material.

It will be found economical in the long run to protect all
exposed brickwork of furnaces or kettles with sheet iron.

In the construction of the refinery building I should advise
brick walls except at the end or side, where there is the greatest
likelihood of future extension; here corrugated iron may be used.
The roof should not be made of corrugated iron, as condensed
or leakage water is liable to collect and drop on those places
where water should be scrupulously avoided. The presence of
water in a mold at the time of casting, even though small in
amount, will cause explosions and will scatter the molten lead,
endangering the workmen.

The item of repair for the ordinary corrugated iron roof may
be diminished by constructing it of 1-in. boards with intervening
spaces of half an inch, the whole overlaid with tarred felt, and
covered with sheets of iron at least No. 27 B. W. G., painted with
graphite paint and joined together with parallel rows of ribbed
crimped iron.

General Expenses. — This item is generally constant, and
calls for no special comment.

Interest. — This important item is, as a rule, considerable,
as the stock of bullion and other gold- and silver-bearing material
is quite large. For this reason special attention should be given
to prevent the accumulation of stock or by-products. The occa-
sional necessity of additional capital to run the business should
preferably be met by an increase of working capital, rather than
by a direct loan.

Expressage. — This item, as a rule, is large, and should be
taken into consideration in the original plans for the location of
the refining works.

Parting. — The item of parting and brokerage is the largest
of the refinery costs, and for obvious reasons a modern smelting
plant should have a parting plant under its own control.

The Working of the By-Products. — This constitutes a large item

102 LEAD SMELTING AND REFINING

of cost, and considerable attention should be devoted to the
improvement of present methods, which seem faulty, slow and
expensive.

Summary. — The items of smaller cost with their respective
amounts per ton of base bullion treated are: Spelter, $0.85; coal,
$0.50; coke, $0.50; supplies, repairs and general expenses, $0.35;
total, $2.10. It is doubtful whether much improvement can be
made in the reduction of these costs.

The items of larger cost are: Labor, $2; interest, $1.32; expres-
sage, $1.10; parting and brokerage, $2; reworking by-products,
$1.50; total, $7.92. The general manager usually attends to the
items of interest, expressage and brokerage, leaving the questions
of labor and working of by-products to the metallurgist.

The cost quoted for smelting practice, as employed at Denver,
will differ necessarily from those at other localities, where the
cost of labor, freight rates on spelter, fuel, etc., are changed.
Refining can doubtless be done at a lower cost at points along
the Mississippi River, and even more so at cities on the Atlantic
seaboard, as Newark or Perth Amboy, N. J.

The consolidation of many of the more important smelting
plants of the United States under one management will doubtless
alter the figures of cost given above, particularly as the interest
cost there stated is at the high rate of 10 per cent., a condition of
affairs now changed to 5 per cent. Other factors have lessened
the cost of refining; the bullion produced at the present time is
softer, or contains a smaller amount of impurities, and admits
of easier working with shorter time and less labor. By proper
management larger tonnages are turned out per man, and the
Howard stirrer and Howard press have simplified and cheapened
the working of the zinc skimmings. To illustrate the compara-
tively recent conditions of cost I have compiled the following
table for each month of the year 1898:

COST OF REFINING DURING 1898, INCLUDING LABOR, SPEL-
TER, COAL, COKE, SUPPLIES, REPAIRS AND GENERAL
EXPENSES.

January $3.59

February 3.28

March 3.26

April 3.59

May $3.38

June 3.56

July 3.65

August 3.54

September $3.35

October 3.45

November 3.20

December. . . . 3 56

Average cost during the year, $3.45.

SMELTING IN THE BLAST FURNACE 103

It is understood, of course, that these figures do not include
cost of interest, expressage, parting, brokerage and reworking
of by-products.

[Although this article refers to conditions in 1898, since which time there
have been improvements in practice, the latter have not been of radical character
and the figures given are fairly representative of present conditions. — EDITOR.]

SMELTING ZINC RETORT RESIDUES1

BY E. M. JOHNSON

(March 22, 1906)

The following notes were taken from work done at the Cherokee
Lanyon Smelter Company, Gas, Kansas, in 1903. It was prac-
tically an experiment. The furnace was only 36 x 90 in. at the
crucible, with a 10-in. side bosh and a 6-in. end bosh. There
were five tuyeres on each side with a 3-in. opening. The side
jackets measured 4.5 ft. x 18 in. The distance from top of crucible
to center of tuyeres was 11.5 in.

The blast was furnished by one No. 4J Connellsville blower.
The furnace originally was only 11 ft. from the center of tuyeres
to the feed-floor, and had only been saving about 60 per cent, of
the lead. This loss of lead, however, was not entirely due to the
low furnace. As no provision had been made to separate the slag
and matte, upon assuming charge I raised the feed-floor 3 ft.,
thereby changing the distance from the tuyere to top of furnace
from 11 ft. to 14 ft. Matte settlers were also installed. These
two changes raised the percentage of lead saved to 92, as shown
by monthly statements. The furnace being small, and a high
percentage of zinc oxide on the charge, the campaigns were
naturally short. The longest run was about six weeks. This
was made on some residue that had been screened from the coarse
coal, and coke, and had weathered for several months. This
particular residue also carried about 10 per cent. lead. The
more recent residue that had not been screened and weathered,
and was low in lead, did not work so well. Although these resi-
dues consisted of a large proportion of coal and coke, it seemed
impossible to reduce the percentage of good lump coke on the
charge lower than 12.5 or 13 per cent. At the same time the
reducing power of the residue was strong, and with the normal
amount of coke caused some trouble in the crucible.

Abstract of a paper in Western Chemist and Metallurgist, I, VII, Aug.,
1905.

104

SMELTING IN THE BLAST FURNACE

105

When residue containing semi-anthracite coal was smelted,
the saving in lead dropped, and the fire went to the top of the
furnace, burning with a blue flame, thereby necessitating the re-
duction of this class of material. This residue had been screened
through a five-mesh screen, and wet down in layers, becoming so
hard that it had to be blasted. The low saving of lead with this
class of material was a surprise, as it has been claimed that the
substitution of part of the fuel by anthracite coal did not affect
the metallurgical operations of the furnace.

The slag was quite liquid and flowed very well at all times.
However, there was a marked variation in the amount at different
tappings. This, I am satisfied, was not due to irregular work on
the furnace, but may be accounted for in the following manner.
The residue (not screened or weathered to any extent), consisting
approximately of one-half coal and coke, was very bulky, and
while there was about 35 per cent, of it on the charge by weight,
there was over 50 per cent, of it by bulk, not including slag and
coke. In feeding, therefore, it was a difficult matter to mix
the whole of it with the charge. Several different ways of feeding
the furnace were tried. The one giving the most satisfactory
results was to feed nearly all of the residue along the center of
the furnace, in connection with the lime-rock, coarse ore and
coarse iron ore, and the fine and easy smelting ores along the
sides. The slag was spread uniformly over the whole furnace,
while the sides were favored with the coke. The charge would
drop several inches at a time, going down a little faster in the
center than on the sides.

It is possible that a small proportion of the residue in con-
nection with the easy smelting, leady, neutral ore, iron ore and
lime-rock formed the type of slag marked No. 1.

Si02

FeO

MnO

CaO

ZnO

1..

33.7

34.1

1.0

16.5

7.5

0.9

0.7

2....

31.0

36.1

1.2

16.0

9.6

1.3

This being tapped with a good flow of slag, the charge would
drop, bringing a proportionately large amount of residue in the
fusion zone which formed the type of slag marked No. 2. There
was also a marked variation in the slag-shells from different pots.

106

LEAD SMELTING AND REFINING

The above cited irregularities of course exist to a certain extent
in any blast furnace.

AVERAGE ANALYSIS OF MATERIALS SMELTED

NAME

Si02

FeO

CaO

MgO

ZnO

A1203

Fe203

Mo. iron ore

10.0

65.0

Lime rock..

1.5

52.0

Mo. galena .

1.5

2.4

9.5

11.0

74.0

Av. of beds .

50.8

16.2

4.6

3.3

9.1

Residuei....

10.5

38.5

18.0

4.8

2.2

1.0

10.0

0.03

Roasted matte2

9.0

48.0

3.0

10.0

4.0

9.9

3.0

21.0

0.06

Barrings. . . .

18.8

24.4

5.0

14.5

6.0

25.4

13.0

0.07

Coke ash. . .

27.0

14.9

4.5

19.7

31.6

H2O

V.M.

F.C.

Ash

Coke*

2.3

85.7

11.1

0.9

ANALYSIS OF BULLION, SLAG AND MATTE PRODUCED

Feb. . .
March

£?:•:
it:::

Sept. .
Average..

LION-^

SiO2

FeO

MnO

CaO

ZnO

SXLA

TTE

1.5
2.5
3.5
4.6
4.0
3.6
2.3

90.0
93.1
104.3
90.0
78.7
90.8
65.3

1.15
1.63
1.59
1.24
1.00
1.21
2.58

31.2
31.3
29.8
30.0
32.2
31.2
32.0

35.9
37.2
37.7
37.3
37.4
37.1
39.7

1.0
1.0
2.7
2.2
1.0
1.7
0.8

14.5
13.9
13.9
14.1
13.9
13.7
14.1

10.3
11.1
11.4
9.3
9.8
9.6
8.1

0.88
0.71
0.52
0.86
0.50
1.10
0.80

0.98
1.30
1.40
1.10
1.15
1.60
1.30

19.0
21.0
23.0
25.4
21.3
23.1
18.6

0.04
0.06
0.07
0.07
0.03
0.08
0.06

8.7
8.0
7.0
5.1
8.9
9.8
7.6

87.5

1.49

31.1

37.5

1.5

14.1

10.0

0.77

1.26

21.6

0.06

7.8

3.0

MONTHLY RECORD OF FURNACE OPERATIONS

OT 0

•

SAVING

« *

Ac Au PB

Feb....
March..

21
21

42.5

44.8

9.0
9.7

12.0
13.5

30.0
37.0

3.7
4.0

8.0)
9.0]

84.4

83.0

90.3

April...

43.7

9.0

13.5

35.0

4.3

10.0

97.9

70.5

96.6

Sky....

49.4

10.0

13.5

30.0

3.5

6.5

95.6

109.5

88.8

July....

41.0

9.8

12.5

34.0

3.8

6.0

97.9

90.0

92.9

August .

47.0

9.3

13.0

32.0

3.7

6.3

86.2

107.5

87.6

Sept.* . .

51.0

7.3

13.0

30.0

2.8

4.6

92.9

94.0

95.6

Average

45.6

9.1

13.0

32.6

3.7

7.2

90.8

92.4

92.0

1 Much better work is being done at present, smelting the Western zinc
ores, and the residue contains about one-third of the above figure, or 7.5 per
cent, of zinc oxide. The high per cent, of ZnO left in residue was mainly due
to poor roasting.

2 There was also considerable coke used of an inferior grade, made from
Kansas coal.

8 Part of the ZnO in roasted matte came from being roasted in the same
furnace the zinc ore had been roasted in.

4 There was less residue on the charges during this month, which accounts
for the larger tonnage with a lower blast.

SMELTING IN THE BLAST FURNACE 107

I believe that, in smelting residues high in zinc oxide, better
metallurgical results would be obtained by using a dry silicious
ore in connection with a high-grade galena ore, provided the
residue be low in sulphur. This was confirmed to a certain degree
in actual practice, as the furnace worked very well upon increasing
the percentage of Cripple Creek ore on the charge. This would
also seem to indicate that alumina had no bad effect on a zinky

ZINC OXIDE IN SLAGS

BY W. MAYNARD HUTCHINGS

(December 24, 1903)

From time to time, in various articles and letters on metallur-
gical subjects in the Engineering and Mining Journal, the question
of the removal of zinc oxide in slags is referred to, and the question
is raised as to the form in which it is contained in the slags.

I gather that opinion is divided as to whether zinc oxide enters
into the slags as a combined silicate, or whether it is simply
carried into them in a state of mechanical mixture.

For many years I have taken great interest in the composition
of slags, and have studied them microscopically and chemically.
The conclusion to which I have been led as regards zinc oxide
is, that in a not too basic slag it is originally mainly, if not wholly,
taken up as silicate along with the other bases. On one occasion,
one of my furnaces for several days produced a slag in which
beautiful crystals of willemite were very abundant, both free in
cavities and also imbedded throughout the mass of solid slag,
as shown in thin sections under the microscope. In the same
slag was a large amount of magnetite, all of which contained a
considerable proportion of zinc oxide combined with it. Mag-
netite crystals, separated out from the slag and treated with
strong acid, yielded shells of material retaining the form of the
original mineral, rich in zinc oxide; an inter-crystallized zinc-iron
spinel, in fact. I have seen and separated zinc-iron spinels very
rich in zinc oxide from other slags. They have been seen in the
slags at Freiberg; and of course everybody knows the very
interesting paper by Stelzner and Schulze, in which they described
the beautiful formations of spinels and willemite in the walls of
the retorts of zinc works.

I think there is thus good ground for concluding that zinc
oxide is slagged off as combined silicate, and that free oxide
does not exist in slags; though zinc oxide does occur in them
after solidification, combined with other oxides, in forms ranging

108

SMELTING IN THE BLAST FURNACE 109

from a zinkiferous magnetite to a more or less impure zinc-iron,
or zinc-iron-alumina spinel, these minerals having crystallized
out in the earlier stages of cooling.

The microscope showed that the crystals of willemite, men-
tioned above, were the first things to crystallize out from the
molten slag. The main constituent was well-crystallized iron-
olivine-fayalite.

PART V
LIME-ROASTING OF GALENA

THE HUNTINGTON-HEBERLEIN PROCESS

(July 6, 1905)

It is a fact, not generally known, that the American Smelting
and Refining Company is now preparing to introduce the Hun-
tington-Heberlein process in all its plants, this action being the
outcome of extensive experimentation with the process. It is
contemplated to employ the process not only for the desulphuri-
zation of all classes of lead ore, but also of mattes. This is a
tardy recognition of the value of a process which has been before
the metallurgical profession for nine years, the British patent
having been issued under date of April 16, 1896, and has already
attained important use in several foreign countries; but it will
be the grandest application in point of magnitude.

The Huntington-Heberlein is the first of a new series of
processes which effect the desulphurization of galena on an entirely
new principle and at great advantage over the old method of
roasting. They act at a comparatively low temperature, so that
the loss of lead and silver is reduced to insignificant proportion;
they eliminate the sulphur to a greater degree; and they deliver
the ore in the form of a cinder, which greatly increases the smelting
speed of the blast furnace. They constitute one of the most
important advances -in the metallurgy of lead. The roasting
process has been the one in which least progress has been made,
and it has remained a costly and wasteful step in the treatment
of sulphide ores. In reducing upward of 2,500,000 tons of ore
per annum, the American Smelting and Refining Company is
obliged to roast upward of 1,000,000 tons of ore and matte.

The Huntington-Heberlein process was invented and first
applied at Pertusola, Italy. It has since been introduced in
Germany, Spain, Great Britain, Mexico, British Columbia, Tas-
mania, and Australia, in the last at the Port Pirie works of the
Broken Hill Proprietary Company. Efforts were made to intro-
duce it in the United States at least five years ago, without success
and with little encouragement. The only share in this metallur-
gical improvement that this country can claim is that Thomas
Huntington, one of the inventors, is an American citizen, Ferdi-
nand Heberlein, the other, being a German.

113

LIME-ROASTING OF GALENA

(September 22, 1905)

The article of Professor Borchers (see p. 116) is, we believe,
the first critical discussion of the reactions involved in the new
methods of desulphurizing galena, as exemplified in the pro-
cesses of Huntington and Heberlein, Savelsberg, and Carmichael
and Bradford, although the subject has been touched upon by
Donald Clark, writing in the Engineering and Mining Journal.
It is perfectly obvious from a study of the metallurgy of these
processes that they introduce an entirely new principle in the
oxidation of galena, as Professor Borchers points out. Inasmuch
as there are already three of these processes and are likely to
be more, it will be necessary to have a type-name for this new
branch of lead metallurgy. We venture to suggest that it may
be referred to as the " lime-roasting of galena," inasmuch as lime
is evidently a requisite hi the process; or, at all events, it is the
agent which will be commonly employed.

When the Huntington-Heberlein process was first described,
it did not even appear a simplification of the ordinary roasting
process, but rather a complication of it. The process attracted
comparatively little attention, and was indeed regarded somewhat
with suspicion. This was largely due to the policy of the com-
pany which acquired the patent rights in refusing to publish the
technical information concerning it that the metallurgical pro-
fession expected and needed. The history of this exploitation is
another example of the disadvantage of secrecy in such matters.
The Huntington-Heberlein process has only become thoroughly
established as a new and valuable departure in metallurgy, a
departure which is indeed revolutionary, nine years after the
date of the original patent. In proprietary processes time is a
particularly valuable element, inasmuch as the life of a patent is
limited.

From the outset the explanation of Huntington and Heberlein
as to the reactions involved in their process was unsatisfactory.
Professor Borchers points out clearly that their conception of

114

LIME-ROASTING OF GALENA 115

the formation of calcium peroxide was erroneous, and indicates
strongly the probability that the active agent is calcium plumbate.
It is very much to be regretted that he did not go further with
his experiments on this subject, and it is to be hoped that they
will be taken up by the professors of metallurgy in other metal-
lurgical schools. The formation of calcium plumbate in the
process was clearly forecasted, however, by Carmichael and
Bradford in their first patent specification; indeed, they con-
sidered that the sintered product consisted largely of calcium
plumbate.

Even yet, we have only a vague idea of the reactions that
occur in these processes. There is undoubtedly a formation of
calcium sulphate, as pointed out by Borchers and Savelsberg;
but that compound is eventually decomposed, since it is one of
the advantages of the lime-roasting that the sintered product is
comparatively low in sulphur. Is it true, however, that the cal-
cium eventually becomes silicate? If so, under what conditions
is calcium silicate formed? The temperature maintained through-
out the process is low, considerably lower than that required for
the formation of any calcium silicate by fusion.

Moreover, it is not only galena which is decomposed by the
new method, but also blende, pyrite and copper sulphides. The
process is employed very successfully in the treatment of Broken
Hill ore that is rather high in zinc sulphide, and it is also to be
employed for the desulphurization of mattes. What are the
reactions that affect the desulphurization of the sulphides other
than lead?

There is a wide field for experimental metallurgy in connection
with these new processes. The important practical development
is that they do actually effect a great economy in the reduction
of lead sulphide ores.

THE NEW METHODS OF DESULPHURIZING GALENA 1
BY W. BORCHERS

(September 2, 1905)

An important revolution in the methods of smelting lead ore,
which had to a large extent remained for centuries unchanged in
their essentials, was wrought by the invention of Huntington and
Heberlein in 1896. More especially is this true of the roast-
reduction method of treating galena, which consists of oxidizing
roasting in a reverberatory furnace and subsequent smelting of
the roasted product in a shaft furnace.

The first stage of the roast-reduction process, as carried out
according to the old method, viz., the oxidizing roast of the galena,
serves to convert the lead sulphide into lead oxide:

PbS + 30 = PbO + S02.

Owing to the basic character of the lead oxide, the production
of a considerable quantity of lead sulphate was of course unavoid-

able:

PbO + SO2 + O = PbSO4.

As this lead sulphate is converted back into sulphide in the
blast-furnace operation, and so adds to the formation of matte,
it has always been the aim (in working up ores containing little
or no copper to be concentrated in the matte) to eliminate the
sulphate as completely as possible, by bringing the charge, es-
pecially toward the end of the roasting operation, into a zone of
the furnace wherein the temperature is sufficiently high to effect
decomposition of the sulphate by silica:

PbSO4 + SiO2 = PbSiO3 + S0

But in the usual mode of carrying out the roast in reverberatory

1 Translation of a paper read before the Naturwissenschaftlicher Verein
at Aachen, and published in Metallurgie, 1905, II, i, 1-6.

116

LIME-ROASTING OF GALENA 117

furnaces, the roasting itself on the one hand, and the decomposi-
tion of the sulphates on the other, were effected only incompletely
and with widely varying results.

Little attention has been paid in connection with the roast-
reduction process to the reaction between sulphates and unde-
composed sulphides, which plays so important a part in the
roast-reaction method of lead smelting. As is well known, lead
sulphate reacts with lead sulphide in varying quantities, forming
either metallic lead or lead oxide, or a mixture of both. A small
quantity of lead sulphate reacting with lead sulphide yields under
certain conditions only lead:

PbSO4 + PbS = Pb2 + 2SO2.

Within certain temperature limits this reaction even proceeds
with liberation of heat. In order to encourage it, it is necessary
to create favorable conditions for the formation of considerable
quantities of sulphate right at the beginning of the operation.
This was first achieved by Huntington and Heberlein, but not in
the simplest nor in the most efficient manner. And, indeed, the
inventors were not by any means on the right track as to the
character of their process, so far as the chemical reactions involved
are concerned.

At first sight the Huntington-Heberlein process does not even
appear as a simplification, but rather as a complication, of the
roasting operation. For in place of the roast carried out in one
apparatus and continuously, there are two roasts which have to
be carried out separately and in two different forms of apparatus;
nevertheless, the ultimate results were so favorable that the
whole process is presumably acknowledged, without reservation,
by all smelters as one of the most important advances in lead
smelting.

It is useful to examine in the light of the German patent
specification (No. 95,601 of Feb. 28, 1897) what were the ideas of
its originators regarding the operation of this process and the
reactions leading to such remarkable results. They stated:

"We have made the observation that when powdered lead
sulphide (PbS), mixed with the powdered oxide of an alkaline
earth metal, e.g., calcium oxide, is exposed to the action of air
#t bright red heat (about 700 deg. C.), and is then allowed to

118 LEAD SMELTING AND REFINING

cool without interrupting the supply of air, an oxidizing decom-
position takes place when dark-red heat (about 500 deg. C.) is
reached, sulphurous acid being expelled, and a considerable
amount of heat evolved; if sufficient air is then continuously
passed through the charge, dense vapors of sulphurous acid
escape, and the mixture gradually sinters together to a mass, in
which the lead of the ore is present in the form of lead oxide,
provided the air blast is continued long enough; there is no need
to supply heat in this process — the heat liberated in the reaction
is quite sufficient to keep it up."

The inventors explained the process as follows:
"At a bright-red heat the calcium oxide (CaO) takes up oxygen
from the air supplied, forming calcium peroxide (CaO2), which
latter afterward, in consequence of cooling down to dark-red heat,
again decomposes into monoxide and oxygen; this nascent oxygen
oxidizes a part of the lead sulphide to lead sulphate, which then
reacts with a further quantity of lead sulphide, with evolution
of sulphur dioxide and formation of lead oxide."

Assuming the formation of calcium peroxide (CaO2), the
process leading to the desulphurization would therefore be repre-
sented as follows:

1. at 700° C CaO + O = CaO2

2. at 500° C 4CaO2 + PbS = 4CaO + PbSO4

3. at the melting point PbS -f PbSO4 = 2PbO + 2SO2 (?)

Reactions 1 and 2 combined, assuming the presence of sufficient
oxygen, give:

PbS + 4CaO + 4O = PbS04 + 4CaO.

Now the invention consists in applying the observation
described above to the working up of galena, and other ores con-
taining lead sulphide, for metallic lead; and the essential novelty
of the process therefore consists in passing air through the mass
cooled to a dark-red heat (500 deg. C.).

This feature sharply distinguishes it from other known pro-
cesses. It is true that in previous processes (compare the Tarno-
witz reverberatory-furnace process, the roasting process used at
Munsterbusch near Stolberg, and others) the lead ore was mixed
with limestone or dolomite (which are converted into oxides in

LIME-ROASTING OF GALENA 119

the early stage of the roast) and the heat was alternately raised
and lowered; but in all cases only a surface action of the air was
produced, the air supply being provided simply by the furnace
draft. Passing air through the mass cooled down, as indicated
above, leads to the important economic advantages of reducing
the fuel consumption, the losses of lead, the manual labor (raking)
and the dimensions of the roasting apparatus.

In order to carry out the process of this invention, the pow-
dered ore is intimately mixed with a quantity of alkaline earth
oxide, e.g., calcium oxide, corresponding to its sulphur content;
if the ore already contains alkaline earth, the quantity to be
added is reduced in accordance. The mixture is heated to
bright-red heat (700 deg. C.) in the reverberatory furnace, in a
strongly oxidizing atmosphere, is then allowed to cool down to
dark-red heat (500 deg. C.), also in strongly oxidizing atmosphere,
is transferred to a vessel called the "converter," and atmospheric
air is passed through at a slight pressure (the inventors have
found a blast corresponding to 35 to 40 cm. head of water suit-
able ).J The heat liberated is quite sufficient to keep the charge
at the reaction temperature, but, if desired, hot blast may also
be used. The mixture sinters together, and (while sulphurous
acid gas escapes) it is gradually converted into a mass consisting
of lead oxide, gangue and calcium sulphate, from which the lead
is extracted in the metallic form, by any of the known methods,
in the shaft furnace. The operation is concluded as soon as the
mass, by continued sintering, has become impermeable to the
blast. If the operation is properly conducted, the gas escaping
contains only small quantities of volatile lead compounds, but
on the other hand up to 8 per cent, by volume of sulphur dioxide.
This latter can be collected and further worked up.

"In place of the oxide of an alkaline earth, ferrous oxide
(FeO) or manganous oxide (MnO) may also be used."

According to the reports on the practice of this process which
have been published,2 conical converters of about 1700 mm.
(5 ft. 6 in.) upper diameter and 1500 mm. (5 ft.) depth are used
in Australian works. At a new plant at Port Pirie (Broken Hill
Proprietary Company) converters 2400 mm. (7 ft. 10 in.) in

1 35 to 40 cm. = 13.78 to 15.75 in. = 8 to 9.12 oz. per sq. in.

2 Engineering and Mining Journal, 1904, LXXVIII, p. 630; article by
Donald Clark; reprinted in this work, p. 144.

120 LEAD SMELTING AND REFINING

diameter and 1800 mm. (5 ft. 11 in.) deep have been installed.
These latter will hold a charge of about eight tons. In the lower
part of these converters, at a distance of about 600 mm. (2 ft.)
from the bottom, there is placed an annular perforated plate,
and upon this a short perforated tube, closed above by a plate
having only a limited number of holes.

No details have been published with regard to the European
installations. The general information which the Metallurgische
Gesellschaft l placed at my disposal upon request some years ago,
for use in my lecture courses, was restricted to data regarding
the consumption of fuel and labor in roasting and smelting the
ores, which was figured at about one-third or one-half of the con-
sumption in the former processes, to the demonstration of the
large output of the comparatively small converters, and to the
reduced size of the roasting plant as the result. But the Euro-
pean establishments which introduced this process were bound
by the owners of the patents, notwithstanding the protection
afforded by the patents, to give no information whatever regarding
the process to outsiders, and not to allow any inspection of the
works.

On the other hand, a great deal appeared in technical literature
which was calculated to excite curiosity. Moreover, as professor
of metallurgy, it was my duty to instruct my pupils concerning
this process among others, and it was therefore very gratifying
to me that one of the students in my laboratory took a special
interest in the treatment of lead ore. I gave him opportunity to
install a small converter, in order to carry out the process on a
small scale, and in spite of the slender dimensions of the apparatus
the very first experiments gave a complete success.

However, I could not harmonize the explanation of the process
given by the inventors with the knowledge which I had acquired
in my many years' practical experience in the manufacture of
peroxides. It is clear from the patent specification that in the
roasting operation at 700 deg. C. a compound must be formed
which functions as an excellent oxygen carrier, for on cooling to
500 deg. C. the further oxidation then proceeds to the end not only
without any external application of heat, but even with vigorous
evolution of heat. No more striking instance than this could
be desired by the theorists who have of recent years again become
1 Owner of the patents. — EDITOR.

LIME-ROASTING OF GALENA 121

so enthusiastic over the idea of catalysis. Huntington and
Heberlein regarded calcium peroxide as the oxygen carrier, but
that is a compound which cannot exist at all under the conditions
which obtain in their process. The peroxides of the alkaline
earths are so very sensitive that in preparing them the small
quantities of carbon dioxide and water must be extracted care-
fully from the air, and yet in the process, in an atmosphere
pregnant with carbon dioxide, water, sulphurous acid, etc., cal-
cium peroxide, the most sensitive of the whole group, is supposed
to form! This could not be.

The only compounds known as oxygen carriers, and capable
of existing under the conditions of the process, are calcium
plumbate and plumbite. I have emphasized this point from the
first in my lectures on metallurgy, when dealing with the Hunt-
ington-Heberlein process, and, in point of fact, this assumption
has since been proved to be correct by the work of L. Huppertz,
one of my students.

During my practical activity (1879-1891) I had prepared
barium peroxide and lead peroxide in large quantities on a manu-
facturing scale, the last-mentioned through the intermediate
formation of plumbites and plumbates:

2NaOH + PbO + O = Na2PbO3 + H2O
or:

4NaOH + PbO + O = Na4Pb04 + 2H2O.

An experiment made in this connection showed that calcium
plumbate is formed just as readily from slacked lime and litharge
as the sodium plumbates above. Litharge is an intermediate
product, produced in large quantities in lead works, and must
in any case be brought back into the process. If, then, the
litharge is roasted at a low temperature with slacked lime, the
roasting of the galena could perhaps be entirely avoided by
introducing that ore together with calcium plumbate into the
converter, after the latter had once been heated up. Mr. Huppertz
undertook the further development of this process, but I have no
information on the later experimental results, as he placed him-
self in communication with neighboring lead works for the purpose
of continuing his investigation, and has not since then given me
any precise data. I will therefore confine myself to the statement

122 LEAD SMELTING AND REFINING

that the fundamental idea for the experiments, which Mr. Hup-
pertz undertook at my suggestion, was the following:

To dispense with the roasting of the galena, which is necessary
according to Huntington and Heberlein; in other words, to convert
the galena by direct blast, with the addition of calcium plumbate,
the latter being produced from the litharge which is an unavoida-
ble intermediate product in the metallurgy of lead and silver.
(Borchers, " Elektrometallurgie," 3d edition, 1902-1903, p. 467.)

This alone would, of course, have meant a considerable sim-
plification of the roast, but the problem of the roasting of galena
has been solved in a better way by A. Savelsberg, of Ramsbeck,
Westphalia, who has determined the conditions for directly con-
verting the galena with the addition of limestone and water and
without previous roasting. He has communicated the following
information regarding these conditions:

In order that, in blowing the air through the mixture of ore
and limestone, an alteration of the mixture may not take place
owing to the lighter particles of the limestone being carried away,
it is necessary (quite at variance with the processes in use hitherto,
in which for the sake of economy stress is laid on the precaution
of charging the ore as dry as possible into the apparatus) to add
a considerable quantity of water to the charge before introducing
it into the converter. The water serves this purpose perfectly,
also preventing any change in the mixture of ore and limestone,
which invariably occurs if the ore is used dry. The water,
moreover, exerts a very beneficial action in the process, inasmuch
as it aids materially in the formation and temporary retention of
sulphuric acid, which latter then, by its oxidizing action, greatly
enhances the reaction and consequently the desulphurization of
the ore. Furthermore, the water tends to moderate the temper-
ature in the charge by absorbing heat in its volatilization.

In carrying out the process the converter must not be filled
entirely all at once, but first only in part, additional layers being
charged in gradually in the course of the operation. In this way
a uniform progress of the reaction in the mass is secured.

The following mode of procedure is advantageously adopted:
A small quantity of glowing fuel (coal, coke, etc.) is introduced
into the converter, which is provided at the bottom with a grate
(perforated sheet iron), the grate being first covered with a thin
layer of crushed limestone in order to protect it from the action

LIME-ROASTING OF GALENA 123

of the red-hot coals and ore. Upon this red-hot fuel a uniform
layer of the wetted mixture of crude ore and limestone is placed.
When the surface of the first layer has acquired a uniform red
heat, a fresh layer is charged on, and this is continued, layer by
layer, until the converter is quite full. While the layers are still
being put on, the blast is passed in at quite a low pressure, and
only when the converter is entirely filled is the whole force of
the blast, at a rather greater pressure, turned on. There then
sets in a kind of slag formation, which, however, is preceded by
a very vigorous desulphurization. After the termination of the
process, which can be recognized by the fact that vapors cease to
be evolved, and that the surface of the ore becomes hard, the
converter is tipped over, and the desulphurized mass drops out
as a solid cone of slag, which is then suitably broken up for the
subsequent smelting in the shaft furnace.

Savelsberg explains the reaction of this process as follows:

"1. The particles of limestone act mechanically, gliding in
between the particles of lead ore and separating them from one
another. In this way a premature sintering is prevented, and
the whole mass is rendered loose and porous.

"2. The limestone moderates the reaction temperature pro-
duced in the combustion of the sulphur, so that the fusion of
the galena, the formation of dust and the separation of metallic
lead are avoided, or at least kept within the limits permissible.
The lowering of the temperature of reaction is due partly to the
decomposition of the limestone into caustic lime and carbon
dioxide, in which heat is absorbed, and partly to the consumption
of the quantity of heat which is necessary in the further progress
of the operation for the formation of a slag from the gangue of
the ore and the lead oxide produced.

"3. The limestone gives rise to chemical reactions. By its
decomposition it produces lime, which, at the moment of its
formation, is converted into calcium sulphate at the expense of
the sulphur in the ore. The calcium sulphate at the time of slag
formation is converted into silicate by the silica present, sulphuric
acid being evolved. The limestone therefore assists directly and
forcibly in the desulphurization of the ore, causing the formation
of sulphuric acid at the expense of the sulphur in the ore, the
sulphuric acid then acting as a strong oxidizing agent toward the;
sulphur in the ore."

124

LEAD SMELTING AND REFINING

The most conclusive proof for the correctness of the opinion
which I expressed above, that it is very important to create at
the beginning of the operation the conditions for the formation
of as much sulphate as possible, has been furnished by Carmichael
and Bradford. They recommend that gypsum be added to the
charge in place of limestone. At one of the works of the Broken
Hill Proprietary Company (where their process has been carried
on successfully, and where lead ores very rich in zinc had to be
worked up) the dehydrated gypsum was mixed with an equal
quantity of concentrate and three times the quantity of slime
from the lead ore-dressing plant, as in the table given herewith:

CONTENTS

OF THE
CONCENTRATE

OF THE
CALCIUM
SULPHATE

fe W

§•

Galena

Zinc blende

Pyrites

2 5

Lime

3 5

4 1

Silica

Sulphur t rioxide . .

The charge is mixed, with addition of water, in a suitable
pug-mill. The mass is then, while still wet, broken up into
pieces 50 mm. (2 in.) in diameter, which are then allowed to dry
on a floor in contact with air; in doing so they set hard, owing to
the rehydration of the gypsum.

As in the case of the Savelsberg process, the converters are
heated with a small quantity of coal, are filled with the material
prepared in the manner above described, and the charge is blown,
regulating the blast in such manner that, after the moisture
present has been dissipated, a gas of about 10 per cent. SO2 con-
tent is produced, which is worked up for sulphuric acid in a
system of lead chambers.

The reactions are in this case the same as in the Savelsberg
process, for here also calcium sulphate is formed transitorily,

LIME-ROASTING OF GALENA 125

which, like other sulphates, reacts partly with sulphides, partly
with silica.

Where gypsum is available and cheap, the Carmichael-Brad-
ford process must be given preference; in all other cases unques-
tionably the Savelsberg process is superior, owing to its great
simplicity.

LIME-ROASTING OF GALENA

BY W. MAYNARD HUTCHINGS

(October 21, 1905)

Much interest attaches to the paper by Professor Borchers,
recently presented in the Engineering and Mining Journal
(Sept. 2, 1905) on "New Methods of Desulphurizing Galena,"
together with an editorial on " Lime-Roasting of Galena"; it is a
curious coincidence that the same issue contained also an article
on the "Newer Treatment of Broken Hill Sulphides," in which is
shown the importance of the new methods as a contribution to
actual practice.

For some years it had been a source of surprise to me that
a new process, so interesting and so successful as the Huntington-
Heberlein treatment of sulphide ores, should have received
scarcely any notice or discussion. This lack, however, now
appears to be remedied. The suggestion that the subject should
be discussed in the Journal is good, as is also that of the desig-
nation "Lime-Roasting" for a type-name. Such observations
and experiments on the subject as I have had occasion to record
have, for many years, figured in my note-books under that
heading.

Whatever may be the final results of the later processes, now
before the metallurgical world or still to come, there can be no
doubt whatever that full and exclusive credit must be given to
Huntington and Heberlein, not only for first drawing attention
to the use of lime, but also for working out and introducing
practically the process. It has been a success from the first;
and so far as part of it is concerned, it seems to be an absolute and
fundamental necessity which later inventors can neither better
nor set aside. The other processes, since patented, however
good they may be, are simply grafts on this parent stem.

It is, however, quite certain that Huntington and Heberlein,
in the theoretical explanation of the process, failed to understand
the most important reactions. Their attributing the effect to

126

LIME-ROASTING OF GALENA 127

the formation and action of calcium peroxide affords a sad case
of a priori assumption devoid of any shred of evidence. As
Professor Borchers points out, calcium peroxide, so difficult to
produce and so unstable when formed, is an absolute and absurd
impossibility under the conditions in question. Probably many
rubbed their eyes with astonishment on reading that part of the
patent on its first appearance, and hastened to look up the chem-
ical authorities to refresh their minds, lest something as to the
nature of calcium peroxide might have escaped them.

Fortunately the patent law is such that there was no danger
of a really good and sound invention being invalidated by a wrong
theoretical explanation by its originators. But, nevertheless, it
was a misfortune that the inventors did not understand their
own process. Had they known, they could have added a few
more words to their patent-claims and rendered the Carmichael
patent an impossibility.

Professor Borchers appears to consider that the active agent
in the new process is calcium plumbate. That this compound
may play a part at some stage of the process may be true; this
long ago suggested itself to some others. We may yet expect to
hear that the experiments undertaken by Professor Borchers him-
self, and by others at his instigation (in which calcium plumbate
is separately prepared and then brought into action with lead
sulphide), have given good results. But it does not appear so
far that there is any real proof that calcium plumbate is formed
in the Huntington-Heberlem or other similar processes; and it is
difficult to see at what stage or how it would be produced under
the conditions in question. This is a point which research may
clear up, but it should not be taken for granted at this stage.
Indeed, it seems to me that the results obtained may be fairly
well explained without calling calcium plumbate into play at all.

Of course the action of lime in contact with lead sulphide
excited interest many years before the new process came into
existence. My own attention to it dates back more than a
dozen years before that time (I was in charge of works where I
found the old " Flintshire process" still in use).

Percy pointed out, in his work on lead smelting, that on the
addition of slacked lime to the charge, at certain stages, to "stiffen
it up," the mixture could be seen to "glow" for a time. When
I myself saw this phenomenon, I commenced to make some

128 LEAD SMELTING AND REFINING

observations and experiments. Also (as others probably had
done) , I had observed that charges of lead with calcareous gangue
are roasted more rapidly and better than others, and to an extent
which could not be wholly explained by simple physical action
of the lime present.

Simple experiments made in assay-scorifiers in a muffle, on
lime roasting, are very striking, and I think quite explain a good
part of what takes place up to a certain stage in the processes
now under consideration. I tried them a number of years ago,
on many sorts of ore, and again more recently, when studying
the working of the new patents. For illustration, I will take
one class of ore (Broken Hill concentrate), using a sample assay-
ing: Pb, 58 per cent.; Fe, 3.6 per cent.; S, 14.6 per cent.; Si02,
3 per cent. The ore contained some pyrite. If two scorifiers
are charged, one with the finely powdered ore alone, and one
with the ore intimately mixed with, say, 10 per cent, of pure
lime, and placed side by side just within a muffle at low redness,
the limed charge will soon be seen to "glow." Before the simple
ore charge shows any sign of action, the limed charge rapidly
ignites all over, like so much tinder, and heats up considerably
above the surrounding temperature, at the same time increasing
noticeably in bulk. This lasts for some time, during which
hardly any SO2 passes off. After the violent glowing is over, the
charge continues to calcine quietly, giving off SO2, but is still far
more active than its neighbor. If, finally, the fully roasted
charge is taken out, cooled and rubbed down, it proves to contain
no free lime at all, but large quantities of calcium sulphate can
be dissolved out by boiling in distilled water. For instance, in
one example where weighed quantities were taken of lime and
the ore mentioned, the final roasted material was shown to
contain nearly 23 per cent, of CaS04; the quantity actually
extracted by water was 20.2 per cent. Further tests show that
the insoluble portion still contains calcium sulphate intimately
combined with lead sulphate, but not extractable by water.

There is no doubt that when lead sulphide (or other sulphide)
is heated with lime, with free access of air, the lime is rapidly
and completely converted into sulphate. The strong base, lime,
apparently plays the part of " catalyzer" in the most vigorous
manner, the first SO2 evolved being instantly oxidized and com-
bined with the lime to sulphate, with so strong an evolution of

LIME-ROASTING OF GALENA 129

heat that the operation spreads rapidly and still goes on ener-
getically, even if the scorifier is taken out of the muffle. Also, the
" catalytic" action starts the oxidation of the sulphides at a far
lower temperature than is required when they are roasted alone.

If, in place of lime, we take an equivalent weight of pure
calcium carbonate and intimately mix it with ore, we obtain
just the same action, only it takes a little longer to start it. Once
started, it is almost as vigorous and rapid, and with the same
results. It does not seem correct to assume (as is usually done)
that the carbonate has first to be decomposed by heat, the lime
then coming into action. The reaction commences in so short a
time and while the charge is still so cool, that no appreciable
driving off of CO2 by heat only can have taken place. The
main liberation of the CO2 occurs during the vigorous exothermic
oxidation of the mixture, and is coincident with the conversion
of the CaO into CaSO4.

If, in place of lime or its carbonate, we use a corresponding
quantity of pure calcium sulphate and mix it with the ore, we
see very energetic roasting in this case also, with copious evolu-
tion of sulphur dioxide, only it is much more energetic and rapid
and occurs at a lower temperature than in the case of a companion
charge of ore alone.

It is very easily demonstrated that the CaSO4 in contact with
the still unoxidized ore (whether it has been introduced ready
made or has been formed from lime or limestone added) greatly
assists the further roasting, in acting as a " carrier" and enabling
calcination to take place more rapidly and easily and at a lower
temperature than would otherwise be the case.

The result of these experiments (whether we mix the ore
with CaO, CaCO3, or CaSO4) is that we arrive with great ease
and rapidity at a nearly dead-sweet roast. The lime is converted
into sulphate, and the lead partly to sulphate and partly to
oxide. Two examples out of several, both from the above ore,
gave results as follows:

No. 1 — Roasted with 20 per cent. CaCO3 (= 11.2 per cent.
CaO); sulphide sulphur, 0.02 per cent.; sulphate sulphur, 9.30
per cent.; total sulphur, 9.32 per cent.

No. 8 — Roasted with 27.2 per cent. CaSO4 (=11. per cent.
CaO); sulphide sulphur, 0.05 per cent.; sulphate sulphur, 11.28
per cent.; total sulphur, 11.33 per cent.

130 LEAD SMELTING AND REFINING

If these calcined products are now intimately mixed with
additional silica (in about the proportions used in the Hunting-
ton-Heberlein process) and strongly heated, fritting is brought
about and the sulphur content is reduced by the decomposition
of the sulphates by the silica. Thus, the resultant material of
experiment No. 1, above, when treated in this manner with
strong heat for three hours, was sintered to a mass which was
quite hard and stony when cold, and which contained 6.75 per
cent, of total sulphur. Longer heating drives out more sulphur,
but a very long time is required; in furnaces, and on a large scale,
it is with great difficulty and cost that a product can be obtained
comparable with that which is rapidly and cheaply turned out
from the " converters" of the new process.

To return to the Huntington-Heberlein process, working, for
example, on an ore more or less like the one given above, we
may assume that, during the comparatively short preliminary
roast, the lime is all rapidly converted into CaSO4 and that some
PbSO4 is also formed (but not much, as the mixture to be trans-
ferred from the furnace to the converter requires not less than
6 to 8 per cent, of sulphur to be still present as sulphide, in order
that the following operation may work at its best). As the
blast permeates the mass, oxidation is energetic; no doubt that
CaSO4 here plays a very important part as a carrier of oxygen,
in the same manner as we can see it act on a scorifier or on the
hearth of a furnace.

What the later reactions are does not seem so clear. They
are quite different from those on the scorifier or on the open
hearth of a furnace, and result in the rapid formation (in successive
layers of the mixture, from the bottom upward) of large amounts
of lead oxide, fluxing the silica and other constituents to a more
or less slaggy mass, which decomposes the sulphates and takes
up the CaO into a complex and easily fused silicate. It is true
that, as a whole, the contents of a well- worked converter are
never very hot, but locally (in the regions where the progressive
reaction and decomposition from below upward is going on) the
temperature reached is considerable. This formation of lead
oxide is so pronounced at times that one may see in the final
product considerable quantities of pure uncombined litharge.

When the work is successful, the mass discharged from the
converters is a basic silicate of PbO, CaO, and oxides of other

LIME-ROASTING OF GALENA 131

metals present, and nearly all the sulphates have disappeared.
A large piece of yellow product (which was taken from a well-
worked converter) contained only 1.1 per cent, of total sul-
phur.

It may be that calcium plumbate is formed and plays a part
in these reactions; but its presence would be difficult to prove,
.and its formation and existence during these stages would not
be easy to explain. Neither does it seem necessary, as the whole
thing appears to be capable of explanation without it.

While the mixture in the converter is still dry and loose,
energetic oxidation of the sulphides goes on, with the intervention
of the CaSO4 as a carrier. As soon as the heat rises sufficiently,
fluxing commences in a given layer and sulphates are decomposed.
The liberated sulphuric anhydride, at the locally high temperature
and under the existing conditions, will act with the greatest
possible vigor on the sulphides in the adjacent layers; these layers
will then in their turn flux and act on those above them, till the
whole charge is worked out. The column of ore is of considerable
hight, requiring a blast of 1J lb., or perhaps more, in the larger
converters now used. This pressure of the oxidizing blast (and
of the far more powerfully oxidizing sulphuric anhydride, con-
tinuously being liberated within the mass of ore, locally very hot)
constitutes a totally different set of conditions from those ob-
tained on the hearth of a furnace with the ore in thin layers,
where it is neither so hot nor under any pressure. It is to
these conditions, in which we have the continued intense action
of red-hot sulphuric anhydride under a considerable pressure
(together with the earlier action of the CaS04), that the remark-
able efficiency of the process seems to me to be due.

In the Carmichael process, the preliminary roast is done
away with, CaSO4 being added directly instead of having to be
formed during the operation from CaO and the oxidized sulphur
of the ore. The charge in the converter has to be started by
heat supplied to it, and the work then goes forward on the same
lines as in the Huntington-Heberlein process, so that we may
assume that the reactions are the same and come under the
same explanation.

Carmichael was quick to see what was really an important
part and a correct explanation of the original process. He was
not misled by wrong theory about any mythical calcium peroxide,

132 LEAD SMELTING AND REFINING

and so he obtained his patent for the use of CaSO4 and the dis-
pensing of the roast in a furnace.

This process would always be limited in its application by the
comparative rarity of cheap supplies of gypsum, but it appears
to be a great success at Broken Hill; there it is not only of im-
portance in working the leady ores, but also for making sulphuric
acid for the new treatment of mixed sulphides by the Delprat
and Potter methods. For this purpose, the use of CaSO4 will
have the additional advantage that the mixture to be worked in
the converter will contain not only the sulphur of the ore, but
also that of the added gypsum; on decomposition, it will yield
stronger gases for the lead chambers of the acid plant.

Finally comes the Savelsberg patent, which is the simplest of
all; not only (like the Carmichael process) avoiding the preliminary
roast with its extra plant, but also not requiring the use of ready-
made CaSO4, as it uses raw ore and limestone directly in the
converter. I have no knowledge as to actual results of this
process; and, so far as I am aware, nothing on the subject has
been published. But Professor Borchers evidently has some infor-
mation about it, and regards it as the most successful of the
methods of carrying out the new ideas. On the face of it, there
seems no reason why it should not attain all the results desired,
as the chemical and physical actions of the CaO, and of the
CaSO4 formed from it, should come into play in the same manner
and in the same order as in the original process; as it is carried
out in the identical converter used by Huntington and Heberlein,
the final reactions (as suggested above) will take place under the
same conditions as to continuous decomposition under considerable
heat and pressure, which I regard as the most vital part of the
whole matter.

It is well to emphasize again the fact that the idea, and the
means of obtaining these vital conditions, owe their origination
to Huntington and Heberlein.

THEORETICAL ASPECTS OF LEAD-ORE ROASTING

BY C. GUILLEMAIN
(March 10, 1906)

It is well known that the process of roasting lead ores in
reverberatory furnaces proceeds in various ways according to
the composition of the ore in question. Thus in roasting a
sulphide lead ore rich in silica, one of the reactions is:

PbS + 3O = PbO + SO2.

But this reaction is incomplete, for the gases which pass on in
the furnace are rich in SO2 and in SO3. And so it is found that
whatever lead oxide is formed passes over almost immediately
into lead sulphate, according to the reaction:

PbO + SO2 + O = PbSO4.

This reaction is the chief one which takes place. Whether
the silicious gangue serves as a catalyzer for the sulphur dioxide,
or whether it serves merely to keep the galena open to the action
of the gases, the end result of the roast is usually the formation
of lead sulphate according to the above reaction.

In the case of an ore rich in galena, a slow roast is essential,
for it is desired to have the following reaction take place during
the latter part of the roast:

PbS + 3PbS04 = 4PbO + 4SO2.

Now, if the heating were too rapid, not enough lead sulphate
would be found to react with the unaltered galena. The quick
roasting of a rich ore would result in the early sintering of the
charge, and sintering prevents the further formation of lead
sulphate. Whether this sintering (which takes place so easily
and which is so harmful in the latter part of the process) is due

1 Abstract of a paper in Metallurgie, II, 18, Sept. 22, 1905, p. 433.

133

134 LEAD SMELTING AND REFINING

to the low melting point of the lead sulphide, whether the heat
evolved by the reaction

PbS + 3O = PbO + SO2

is sufficient to melt the lead sulphide, or whether other thermo-
chemical effects (notably the preliminary sulphatizing of the
lead sulphide) come into play, must for the present be undecided.
Suffice it to say that the sintering of the charge works against a
good roast.

In the Tarnowitz process a definite amount of lead sulphide
is converted into lead sulphate by a preliminary roast. The
sulphate then reacts with the unaltered lead sulphide, and metallic
lead is set free, thus:

PbS + PbSO4 = 2Pb + 2SO2.

But when a very little of the sulphide has been transformed
into sulphate, and when there is so little of the latter present that
only a small amount of lead sulphide can be reduced to metallic
lead, the mass of ore begins to sinter and grow pasty. Very little
lead could be formed were it not for the addition of crushed lime
to the charge just before the sintering begins. This lime breaks
up the charge and cools it, prevents any sintering, and allows
the continued formation of lead sulphate.

It scarcely can be held that the lime has any chemical effect
in forming lead sulphate, or in forming a hypothetical compound
of lead and calcium. Even if such theories were tenable from a
physico-chemical point of view, they would be lessened in impor-
tance by the fact that other substances, such as purple ore or
puddle cinder, act just as well as the lime.

There are now to be mentioned several new processes of
lead-ore roasting whose operations fall so far outside the common
ideas on the subject that their investigation is full of interest.
For a long time the attempt had been made to produce lead
directly by blowing air through lead sulphide in a manner analo-
gous to the production of bessemer steel or the converting of
copper matte. In the case of the lead sulphide, the oxidation
of the sulphur was to furnish the heat necessary to carry on the
process.

After many attempts along this line, Antonin Germot has

LIME-ROASTING OF GALENA 135

perfected a method wherein, by blowing air through molten
galena, metallic lead is obtained.1 About 60 per cent, of a pre-
viously melted charge of galena is sublimed as lead sulphide, and
the rest remains behind as metallic lead. The disadvantages of
the process are the difficulties of collecting all of the sublimate
and of working it up. Moreover, it is impossible as yet to secure
two products of which one is silver-free and the other silver-
bearing. The silver values are in both the metallic lead and in
the sublimed lead sulphide.

While the process just described answers for pure galena, it
fails with ores which contain about 10 per cent, of gangue. In
the case of such ores, they form a non-homogeneous mass when
melted, and the blast penetrates the charge with difficulty. If
the pressure is increased the air forces itself out through tubes
and canals which it makes for itself, and the charge freezes around
these passages.

Messrs. Huntington and Heberlein have gone a little farther.
Although they are unable to obtain metallic lead directly, they
prepare the ore satisfactorily for smelting in the blast furnace,
after their roasting is completed. The inventors found that if
lead sulphide is mixed with crushed lime, heated with access of
air, and then charged into a converter and blown, the sulphur is
completely removed in the form of sulphur dioxide. The charge,
being divided by the lime, remains open uniformly to the passage
of air, and sinters only when the sulphur is eliminated.

The inventors announce, as the theory of their process, that
at 700 deg. C. the lime forms a dioxide of calcium (Ca02) which
at 500 deg. C. breaks down into lime (CaO) and nascent oxygen.
This nascent oxygen oxidizes the lead sulphide to lead sulphate
according to the reaction:

PbS + 4O = PbSO4.

Furthermore it is claimed that the heat evolved by this last
reaction is large enough to start and keep in operation a second
reaction, namely

PbS + PbSO4 = 2PbO + 2SO2.

The theory, as just mentioned, cannot be accepted, and some of
the reasons leading to its rejection will be given.

1 This method is described further on in this book.

136 LEAD SMELTING AND REFINING

It is well established that the simple heating of lime with
access of air will not result in further oxidation of the calcium.
The dioxide of calcium cannot be formed even by heating lime
to incandescence in an atmosphere of oxygen, nor by fusing lime
with potassium chlorate. Moreover, calcium stands very near
barium in the periodic system. And as the dioxide of barium
is formed at a low temperature and breaks up on continued
heating, it seems absurd to suppose that the dioxide of calcium
would act in exactly the opposite manner. Moreover, a consid-
eration of the thermo-chemical effects will disclose more incon-
sistencies in the ideas of the inventors. The breaking up of
CaO2 into CaO and O is accompanied by the evolution of 12 cal.
The reaction of the oxygen (thus supposed to be liberated) upon
the lead sulphide is strongly exothermic, giving up 195.4 cal.
So much heat is produced by these two reactions that, if the ideas
of the inventors were true, the further breaking up of the calcium
dioxide would stop, as the whole charge would be above 500
deg. C. It appears, then, that the explanations suggested by
Messrs. Huntington and Heberlein are untrue.

In the usual roasting process, as carried out in reverberatory
furnaces, it is well established that the gangue, and whatever
other substances are added to the ore, prevent mechanical locking
up of charge particles, since they stop sintering. It is not at all
improbable that in the new roasting process the chief, if not the
only, part played by the lime is the same as that played by the
gangue in reverberatory-furnace roasting. A few observations
leading to this belief will be given.

It is known that other substances will answer just as well as
lime in this new roasting process. Such substances are manganese
and iron oxides. Not only these two substances, but in fact any
substance which answers the purpose of diminishing the local
strong evolution of heat, due to the reaction:

PbS + 3O = PbO + SO2,

serves just as well as the lime. This fact is proved by exhaustive
experiments in which mixtures of lead sulphide on the one hand,
and quartz, crushed lead slags, iron slags, crushed iron ores,
crushed copper slags, etc., on the other hand, were used for
blowing. All these substances are such that any chemical action,
analogous to the splitting up of Ca02, or the formation of plum-

LIME-ROASTING OF GALENA 137

bates as suggested by Dr. Borchers, cannot be imagined. The
time is not yet ripe, without more experiments on the subject, to
assert conclusively that there is no acceleration of the process
due to the formation of plumbates through the agency of lime.
But the facts thus far secured point out that such reactions are,
at least, not of much importance.

Theoretical considerations point out that it ought to be
possible to avoid the injurious local increase of temperature
during the progress of this new roasting process, without having
to add any substance whatever. To explain: The first reaction
taking place in the roasting is

PbS + 3O = PbO + SO2 + 99.8 cal.

Now the heat thus liberated may be successfully dispersed if
there is, in simultaneous progress, the endothermic reaction:

PbS + 3PbSO4 = 4PbO + 4SO2 - 187 cal.

Hence if there could be obtained a mixture of lead sulphide
and of lead sulphate in the proportions demanded by the above
reaction, then such a mixture ought to be blown successfully
to lead oxide without the addition of any other substance.
Such a process has, in fact, been carried out. The original
galena is heated until the required amount of lead sulphate has
been formed. Then the mixture of lead sulphide and of lead
sulphate is transferred to a converter and blown successfully
without the addition of any other substance.

The adaptability of an ore to the process just mentioned
depends on the cost of the preliminary roast and the thoroughness
with which it must be done. As is known, when lead sulphide
is heated with access of air, it is very easy to form sintered incrus-
tations of lead sulphate. If these incrustations are not broken
up, or if their formation is not prevented by diligent rabbling,
the further access of air to the mass is prevented and the oxida-
tion of the charge stops. If ores with such incrustations are
placed in the converter without being crushed, they remain
unaltered by the blowing. If the incrustations are too numerous
the converting becomes a failure.

It has been found that the adoption of mechanical roasting
furnaces prevents this. Such furnaces appear to stop the fre-
quent failures of the blowing which are due to the lack of care

138 LEAD SMELTING AND REFINING

on the part of the workmen during the preliminary roasting.
Moreover, in such mechanical furnaces a more intimate mixture
of the sulphide with the sulphate is obtained, and the degree of
the sulphatizing roast is more easily controlled.

As a summary of the facts connected with this new blowing
process, it may be stated that the best method of working can
be determined upon and adopted if one has in mind the fact
that the amount of substance (lime) to be added is dependent
on: 1, the amount of sulphur present; 2, the forms of oxidation
of this sulphur; 3, the amount of gangue in the ore; 4, the specific
heats of the gangue and of the substance added; 5, the degree of
the preparatory roasting and heating.

For example, with concentrates which run high in sulphur,
there is required either a large amount of additional material,
or a long preliminary roast. The specific heat of the added
material must be high, and the heat evolved by the oxidation of
the sulphur in the preliminary roast must be dispersed. Often-
times it is necessary to cool the charge partially with water
before blowing. On the other hand, if the ore runs low in sulphur,
the preliminary roast must be short, and the temperature neces-
sary for starting the blowing reactions must be secured by heating
the charge out of contact with air. Not only must no flux be
added, but oftentimes some other sulphides must be supplied in
order that the blowing may be carried out at all.

The opportunity for the acquisition of more knowledge on
this subject is very great. It lies in the direction of seeing whether
or not the strong local evolution of heat cannot be reduced by
blowing with gases poor in oxygen rather than with air. Mixtures
of filtered flue gases and of air can be made in almost any pro-
portion, and such mixtures would have a marked effect upon the
possibility of regulating the progress of the oxidation of the
various ores and ore-mixtures which are met with in practice.

METALLURGICAL BEHAVIOR OF LEAD SULPHIDE
AND CALCIUM SULPHATE1

BY F. O. DOELTZ

(January 27, 1906)

In his British patent,2 for desulphurizing sulphide ores, A. D.
Carmichael states that a mixture of lead sulphide and calcium
sulphate reacts "at dull red heat, say about 400 deg. C.," forming
lead sulphate and calcium sulphide, according to the equation:

PbS + CaS04 = PbS04 + CaS.

Judging from thermo-chemical data, this reaction does not
seem probable. According to Roberts-Austen,3 the heats of for-
mation (in kilogram-calories) of the different compounds in this
equation are as follows: PbS = 17.8; CaSO4 = 318.4; PbSO4 =
216.2; CaS = 92. Hence we have the algebraic sum:

- 17.8 - 318.4 + 216.2 + 92 = - 28.0 cal.

As the law of maximum work does not hold, experiment only
can decide whether this decomposition takes place or not. The
following experiments were made:

Experiment 1. — Coarsely crystalline and specially pure galena
was ground to powder. Some gypsum was powdered, and then
calcined. The powdered galena and calcined gypsum were mixed
in molecular proportions (PbS -f CaSO4), and heated for 1J hours
to 400 deg. C., in a stream of carbon dioxide in a platinum resist-
ance furnace. The temperature was measured with a Le Chatelier
pyrometer. The material was allowed to cool in a current of
carbon dioxide.

The mixture showed no signs of reaction. Under the magni-
fying glass the bright cube-faces of galena could be clearly dis-

* Translated from Metallurgie, Vol. II, No. 19.

2 British patent, No. 17,580, Jan. 30, 1902, "Improved process for de-
sulphurizing sulphide ores."

3 W. C. Roberts-Austen, "An Introduction to the Study of Metallurgy,'*
London, 1902.

139

140 LEAD SMELTING AND REFINING

tinguished. If any reaction had taken place, in accordance with
the equation given above, no bright faces of galena would have
remained.

Experiment 2. — A similar mixture was slowly heated, also in
the electric furnace, to 850 deg. C., in a stream of carbon dioxide,
and was kept at this temperature for one hour.

It was observed that some galena sublimed without decom-
position, being redeposited at the colder end of the porcelain
boat (7 cm. long), in the form of small shining crystals. The
residue was a mixture of dark particles of galena and white
particles of gypsum, in which no evidence of any reaction was
visible under the microscope. That galena sublimes markedly
below its melting point has already been noted by Lodin.1

Experiment 3. — In order to determine whether the inverse
reaction takes place, for which the heat of reaction is + 28.0 cal.,
the following equations are given:

PbSO4 + CaS = PbS + CaSO4;
- 216.2 - 92 + 17.8 + 318.4 = 28.

A mixture of lead sulphate and calcium sulphide was heated
in a porcelain crucible in a benzine-bunsen flame (Barthel burner) .
The materials were supplied expressly "for scientific investiga-
tion" by the firm, C. A. F. Kahlbaum.

The white mixture turned dark and presently assumed the
color which would correspond to its conversion into lead sulphide
and calcium sulphate. This experiment is easy to perform.

Experiment 4. — The same materials, lead sulphate and cal-
cium sulphide, were mixed in molecular ratio (PbSO4 -f CaS),
and were heated for 30 minutes to 400 deg. C., on a porcelain
boat in the electric furnace, in a current of carbon dioxide. The
mixture was allowed to cool in a stream of carbon dioxide, and
was withdrawn from the furnace the next day (the experiment
having been made in the evening).

The mixture showed a dark coloration, similar to that of the
last experiment; but a few white particles were still recognizable.
The material in the boat smelled of hydrogen sulphide.

Experiment 5. — A mixture of pure galena and calcined
gypsum, in molecular ratio (PbS -f CaSO4), was placed on a

1 A. Lodin, Comptes rendus, 1895, CXX, 1164-1167; Berg. u. Hiittenm.
Ztg., 1903, p. 63.

LIMi^ROASTING OF GALENA 141

covered scorifier and introduced into the hot muffle of a petroleum
furnace, at 700 to 800 deg. C. The temperature was then raised
to 1100 deg. C.

From 5 g. of the mixture a dark-gray porous cake weighing
3.7 was thus obtained. There was some undecomposed gypsum
present, recognizable under the magnifying glass. No metallic
lead had separated out. When hot hydrochloric acid was poured
over the mixture, it evolved hydrogen sulphide. The fracture
of the cake showed isolated shining spots. The supposition that
it was melted or sublimed galena was confirmed by the aspect of
the cake when cut with a knife; the surface showed the typical
appearance of the cut surface of melted galena. On cutting, the
cake was found to be brittle, with a tendency to crumble. On
boiling with acetic acid, a little lead went into solution. Wetting
with water did not change the color of the crushed cake.

Experiment 6. — In his experiments for determining the
melting point of galena, Lodin l found that, in addition to its
sublimation at a comparatively low temperature, the galena also
undergoes oxidation if carbon dioxide is used as the "neutral"
atmosphere. Lodin was therefore compelled to use a stream of
nitrogen in his determination of the melting point of galena.
Now the temperature of experiment 2 (850 deg. C.), described
heretofore, is not as high as the melting point of galena (which
lies between 930 and 940 deg. C.); therefore experiment 2 was
repeated in a stream of nitrogen, so as to insure a really neutral
atmosphere. A mixture of galena and calcined gypsum in mole-
cular ratio (PbS -f CaSO4) was heated to 850 deg. C., was kept at
this temperature for one hour, and allowed to cool, the entire
operation being carried out in a stream of nitrogen.

Again, galena had sublimed away from the hotter end of the
porcelain boat (6.5 cm. long), and had been partially deposited
in the form of small crystals of lead sulphide at the colder end.
The material in the boat consisted of a mixture of particles having
the dark color of galena, and others with the white color of gyp-
sum, the original crystals of gypsum and the bright surfaces of
the lead sulphide being distinctly recognizable under the magni-
fying glass. The loss in weight was 1.9 per cent.

Experiment 7. — For the same reason as in 2, experiment 5
was also repeated, using a current of nitrogen. A mixture oi
1 Comptes rendus, loc. cit.

142 LEAD SMELTING AND REFINING

galena and calcined gypsum, in molecular ratio (PbS -f CaS04)
was heated in a porcelain boat to 1030 deg. C., in a platinum-
resistance furnace, and allowed to cool, being surrounded by a
stream of nitrogen during the whole period.

Some sublimation of lead sulphide again took place. The
mixture was seen to consist of white particles of gypsum, and
others dark, like galena. The loss in weight was 3.5 per cent.
The mixture had sintered together slightly ; with hot hydrochloric
acid, it evolved hydrogen sulphide. On boiling with acetic acid,
a little lead (only a trace) went into solution. There was,
therefore, practically no lead oxide present; no metallic lead had
separated out.

Experiment 8. — In experiment 3, lead sulphate and calcium
sulphide were mixed roughly and by hand (i.e., not weighed out
in molecular ratio); in this experiment such a mixture of lead
sulphate and calcium sulphide in molecular ratio (PbSO4 + CaS)
was heated in a porcelain crucible in a benzine-bunsen flame.
It presently turned dark, and a dark gray product was obtained,
as in the former experiment.

Experiment 9. — In a mixture of lead sulphate and sodium
sulphide in molecular ratio (PbSO4 -f Na2S), the constituents
react directly on rubbing together in a porcelain mortar. The
mass turns dark gray, with formation of lead sulphide and sodium
sulphate.

If a similar mixture is heated, it also turns dark gray. On
lixiviation with water, a solution is obtained which gives a dense
white precipitate with barium chloride.

Experiment 10. — If lead sulphate and calcium sulphide are
rubbed together in a mortar, the mass turns a grayish-black.

Conclusion. — From these experiments I infer that the
reaction

PbS + CaS04 = PbSO4 + CaS

does not take place, but, on the contrary, that when lead sulphate
and calcium sulphide are brought together, the tendency is to
form lead sulphide and calcium sulphate.

Nevertheless, on heating a mixture of galena and gypsum in
contact with ah-, lead sulphate will be formed along with lead
oxide; not, however, owing to any double decomposition of the
galena with the gypsum, but rather to the formation of lead

LIME-ROASTING OF GALENA 143

sulphate from lead oxide and sulphuric acid produced by catalysis,
thus:

PbO + S02 + O = PbS04.

This is the well-known process which always takes place in
roasting galena, the explanation of which was familiar to Carl
Friedrich Plattner. That the presence of gypsum has any
chemical influence on this process seems to be out of the question
according to the above experiments.

THE HUNTINGTON-HEBERLEIN PROCESS

BY DONALD CLARK

(October 20, 1904)

The process was patented in 1897, and is based on the fact
that galena can be desulphurized by mixing it with lime and
blowing a current of air through the mixture. If the temperature
is dull red at the start, no additional source of heat is necessary,
because the reaction causes a great rise in temperature. The
chemistry of the process cannot be said at present to have been
worked out in detail.

The reactions given by the patentees are not satisfactory, since
calcium dioxide is formed only at low temperatures and is readily
decomposed on gently warming it; lead oxide, however, combines
with oxygen under suitable conditions at a temperature not
exceeding 450 deg. C. and forms a higher oxide, and it is probable
that this unites with the lime to form calcium plumbate. The
reaction between sulphides and lime when intimately mixed and
heated may be put down as

CaO + PbS = CaS + PbO.

In contact with the air the calcium sulphide oxidizes to sulphite,
then to sulphate, then reacts with lead oxide, giving calcium
plumbate and sulphur dioxide,

CaS04 + PbO = CaPbO3 + SO2.

Further, calcium sulphate will also react with galena, giving
calcium sulphide and lead sulphate; the calcium sulphide is oxi-
dized, by air blown through, to calcium sulphate again, the
ultimate reaction being

CaSO4 4- PbS + O = CaPbO3 + SO2.

In all cases the action is oxidizing and desulphurizing. It was

144

LIME-ROASTING OF GALENA 145

found that oxides of iron and manganese will, to a certain extent,
serve the same purpose as lime, and on application to complex
ores, especially those containing much blende, that these may
be desulphurized as well as galena. In the case of zinc sulphide
the decomposition is probably due to the interaction of sulphide
and sulphate.

ZnS + 3ZnSO4 = 4ZnO + 4SO2.

The process has now been adopted by the Broken Hill Proprie-
tary Company at its works at Port Pirie, the Tasmanian Smelting
Company, Zeehan, the Fremantle Smelting Works, West Aus-
tralia, and the Sulphide Corporation's works at Cockle Creek,
New South Wales.

The operations carried on at the Tasmania Smelting Works
comprise mixing pulverized limestone, galena and slag-making
materials and introducing the mixture either into hand-rabbled
reverberatories or mechanical furnaces with rotating hearths.
After a roast, during which the materials have become well
mixed and most of the limestone converted into sulphate and
about half of the sulphur expelled, the granular product is run
while still hot into the Huntington-Heberlein converters. These
consist of inverted sheet-iron cones, hung on trunnions, the
diameter being 5 ft. 6 in. and the depth 5 ft. A perforated plate
or colander is placed as a diaphragm across the apex of the cone,
the small conical space below serving as a wind-box into which
compressed air is forced. A hood above the converter serves to
carry away waste gases. As soon as the vessel is rilled, air under
a pressure of 17 oz. is forced through the mass, which rapidly
warms up, giving off sulphur dioxide abundantly. The tempera-
ture rises and the mixture fuses, and in from two to four hours
the action is complete. The sulphur is reduced from 10 to
1 per cent., and the whole mass is fritted and fused together.
The converter is emptied by inverting it, when the sintered mass
falls out and is broken up and sent to the smelters. There are
12 converters, of the size indicated, for the two mechanical
furnaces, of 15 ft. diameter. Larger converters of the same
type were erected to deal with the product from the hand-rabbled
roasters.

At Cockle Creek, New South Wales, the galena concentrate
is reduced to 1.5 mm., more than 60 per cent, of the material

146 LEAD SMELTING AND REFINING

being finer; the limestone is crushed down to from 10 to 16 mesh;
silica is also added, if it does not exist in the ore, so that, excluding
the lead, the rest of the bases will be in such proportion as to
form a slag running about 20 per cent, silica. The mixture may
contain from 25 to 50 per cent, lead, and from 6 to 9 per cent,
lime; if too much lime is added the final product is powdery,
instead of being in a fused condition. This is given a preliminary
roast in a Godfrey furnace.

The Godfrey furnace is characterized by a rotating, circular
hearth and a low dome-shaped roof. Ore is fed through a
hopper at the center and deflected outward by blades
attached to a fixed radial arm. At each revolution the ore
is turned over and moved outward, the mount of deflection of
the blades, which are adjustable, and rate of rotation of the
hearth, determining the output.

The hot semi-roasted ore is discharged through a slot at
the circumference of the roaster. This may contain from 12 to
6.5 per cent, of sulphur, but from 6.5 to 8 per cent, is held to be
the most suitable quantity for the subsequent operations. Thor-
ough mixing is of the utmost importance, for if this is not done
the mass will ''volcano" in the converter; that is, channels will
form in the mass through which the gases will escape, leaving
lumps of untouched material alongside. The action can be
started if a little red-hot ore is run into the converter and cold
ore placed above it; the whole mass will become heated up, and
the products will fuse, and sinter into a homogeneous mass
showing none of the original ingredients. At Cockle Creek the
time taken is stated to be five hours; a small air-pressure is turned
on at first, and ultimately it is increased to 20 oz.

Operations at Port Pirie are conducted on a much larger
scale. A mixture of pulverized galena, powdery limestone, iron-
stone and sand is fed into Ropp furnaces, of which there are five,
by means of a fluted roll placed at the base of a hopper. Each
roaster deals with 100 tons of the mixture in 24 hours. About
50 per cent, of the sulphur is eliminated from the ore by the
Ropps (the galena in this case being admixed with a large amount
of blende, there being only 55 per cent, of lead and 10 per cent,
of zinc in the concentrate produced at the Proprietary mine).
The hot ore from the roasters is trucked to the converters, there
being 17 of these ranged in line. The converters here are large

LIME-ROASTING OF GALENA 147

segmental cast-iron pots hung on trunnions; each is about 8 ft.
diameter and 6 ft. deep, and holds an 8-ton charge. At about
two feet from the bottom an annular perforated plate fits hori-
zontally; a shallow frustrum of a cone, also perforated, rests on
this; while a plate with a few perforations closes the top of the
frustrum. The whole serves as a wind-box. A conical hood
with flanged edges rests on the flanged edges of the converter,
giving a close joint. This hood is provided with doors which
allow the charge to be barred if necessary. A pipe about 1 ft.
9 in. diameter, fitted with a telescopic sliding arrangement, allows
for the raising or lowering of the hood by block and tackle, and
thus enables the converter to be tilted up and its products emptied.
The cast-iron pots stand very well; they crack sometimes, but
they can be patched up with an iron strap and rivets. Only two
pots have been lost in 18 months.

Air enters at a pressure of about 24 oz. and the time taken
for conversion is about four hours. The sulphur contents are
reduced to about three per cent. It is found that the top of the
charge is not so well converted as the interior. There is prac-
tically no loss of lead or silver due to volatilization and very
little due to escape of zinc. It has also been found that practically
all the limestone fed into the Ropp is converted into calcium
sulphate; also that a considerable portion of lead becomes sul-
phate, and it is considered that lead sulphate is as necessary for
the process as galena.

The value of the process may be judged from the fact that
better work is now done with 8 blast furnaces than was done
with 13 before the process was adopted. In addition to the
sintered product from the Huntington-Heberlein pots, sintered
slime, obtained by heap roasting, and flux consisting of limestone
and ironstone, are fed into the furnaces, which take 2000 long
tons per day of ore, fluxes and fuel. The slags now being pro-
duced average: Si02, 25 to 26 per cent.; FeO, 1 to 3 per cent.;
MnO, 5 to 5.5; CaO, 15.5 to 17; ZnO, 13; A12O3> 6.5; S, 3 to 4;
Pb, by wet assay, 1.2 to 1.5 per cent.; and Ag, 0.7 oz. per ton.
Although this comparatively large quantity of sulphur remains,
yet no matte is formed.

THE HUNTINGTON-HEBERLEIN PROCESS AT
FRIEDRICHSHUTTE l

BY A. BlERNBAUM
(September 2, 1905)

Nothing, for some time past, has caused such a stir in the
metallurgical treatment of lead ores, and produced such radical
changes at many lead smelting works, as the introduction of the
Huntington-Heberlein process. This process (which it may be
remarked, incidentally, has given rise to the invention of several
similar processes) represents an important advance in lead
smelting, and, now that it has been in use for some time at the
Friedrichshutte, near Tarnowitz, in Upper Silesia, and has there
undergone further improvement in several respects, a comparison
of this process with the earlier roasting process is of interest.

At the above-mentioned works, up to 1900 the lead ore was
treated exclusively (1) by smelting in reverberatory furnaces
(Tarnowitzerof en) , and (2) by roasting in reverberatory sintering-
furnaces (Fortschaufelungsofen), with subsequent smelting of the
roasted material in the shaft furnace. The factor which deter-
mined whether the treatment was to be effected in the reverber-
atory-smelting or in the roasting-sintering furnace was the per-
centage of lead and zinc in the ores; those comparatively rich in
lead and poor in zinc being worked up in the former, with partial
production of pig-lead; while those poorer in lead and richer in
zinc were treated in the latter. About two-fifths of the lead ores
annually worked up were charged into the reverberatory-smelting
furnaces, and three-fifths into the sintering furnaces.

In 1900 there were available 10 reverberatory-smelting and
nine sintering furnaces. These were worked exclusively by hand.

The sintered product of the roasting furnaces, and the gray
slag from the reverberatory-smelting furnaces, were transferred
to the shaft furnaces for further treatment, and were therein

1 Translated from the Zeitschrift filr das Berg.- Hiitten- und Salinenwesen
im. preuss. Staate, 1905, LIII, ii, pp. 219-230.

148

LIME-ROASTING OF GALENA 149

smelted together with the requisite fluxes. Eight such furnaces
(8 m. high, and 1.4 m., 1.6 m., and 1.8 m. respectively in diameter
at the tuyeres), partly with three and partly with five or eight
tuyeres, were at that time in use.

Now that the Huntington-Heberlein process has been com-
pletely installed, the reverberatory-smelting furnaces have been
shut down entirely, and the sintering furnaces also for the most
part; all kinds of lead ore, with a single exception, are worked up
by the Huntington-Heberlein process, irrespective of the contents
of lead and zinc. An exceedingly small proportion of the ore
treated, viz., the low-grade concentrate (Herdschlieche) containing
25 to 35 per cent. Pb, is still roasted in the old sintering furnace,
together with various between-products (such as dust, fume,
scaffoldings, and matte); these are scorified by the aid of the high
percentage of silica in the material.

For roasting lead ores at the present time there are six round
mechanical roasters of 6-m. diameter, one of 8-m. diameter, and
two ordinary, stationary Huntington-Heberlein furnaces. The
latter (which represent the primitive Huntington-Heberlein fur-
naces, requiring manual labor) have recently been shut down,
and will probably never be used again. In the mechanical
Huntington-Heberlein furnace, roasting of lead ore is carried only
to such a point that a small portion of the lead sulphide is con-
verted into sulphate. The desulphurization of the ore is com-
pleted in the so-called converter (made of iron, pear-shaped or
hemispherical in form) in which the charge, up to this stage
loosely mixed, is blown to a solid mass.

Owing to the ready fusibility of this product (which still
contains, as a rule, up to 1.5 per cent, sulphur as sulphide), it is
possible to use shaft furnaces of rather large dimensions ; therefore
a round shaft furnace (2.4 m. diameter at the tuyeres, 7 m. high,
and furnished with 15 tuyeres) was built. In this furnace nearly
the whole of the roasted ore from the Huntington-Heberlein
converters is now smelted, some of the smaller shaft furnaces
being used occasionally. The introduction of the new process
has caused no noteworthy change in the subsequent treatment
of the work-lead.

In the following study I shall discuss the treatment of a given
annual quantity of ore (50,000 tons), which is the actual figure
at the Friedrichshutte at the present time.

150 LEAD SMELTING AND REFINING

1 . Roasting Furnaces. — A reverberatory-smelting furnace
used to treat 5 tons of ore in 24 hours; a roasting-sintering fur-
nace, 8 tons. Assuming the ratios previously stated, the annual
treatment by the former process would be 20,000 tons, and by
the latter 30,000 tons. On the basis of 300 working days per
year, and no prolonged stoppages for furnace repairs (though
considering the high temperatures of these furnaces this record
would hardly be expected), there would be required:

20,000 -T- (5 X 300) = 13.3 (or 13 to 14 reverberatory furnaces).
30,000 -M8 X 300) = 12.5 (or 12 to 13 sintering furnaces).

The capacity of a stationary Huntington-Heberlein furnace is
18 tons; hence in order to treat the same quantity of ores there
would be required:

50,000 -f- (18 X 300) = 9.3 (or 9 to 10 Huntington-Heberlein furnaces).

With the revolving-hearth roasters (of 6 m. diameter) working
a total charge of at least 27 tons of ore, there would be required :

50,000 •*- (27 X 300) = 6.1 (or 6 to 7 roasters).

Still better results are obtained with the 8-m. round roaster,
which has been in operation for some time; in this, 55 tons of ore
can be roasted daily. Three such furnaces would therefore suffice
for working up the whole of the ore charged per annum.

Now, making due provision for reserve furnaces, to work up
50,000 tons of ore would require:

Reverberatory (15) and sintering furnaces (15) 30

Stationary Huntington-Heberlein furnaces 12

6-m. revolving-hearth furnaces 8

8-m. revolving-hearth furnaces 4

Similar relations hold good regarding the number of workmen
attending the furnaces, there being required, daily, six men for
the reverberatory furnace; eight men for the sintering furnace;
ten men for the stationary; and six men for the mechanical
Huntington-Heberlein furnace; or, for 14 reverberatory furnaces,
daily, 84 men; for sintering furnaces, daily, 104 men; total,
188 men. While for 10 stationary Huntington-Heberlein furnaces,
100 men are required ; and for 7 mechanical Huntington-Heberlein
furnaces, daily, 42 men. It is expected that only 14 men (working

LIME-ROASTING OF GALENA 151

in two shifts) will be required to run the new installation with
8-m. round roasters.

It is true that the exclusion of human labor here has been
carried to an extreme. The roasters and converters will be
charged exclusively by mechanical means; thus every contact of
the workmen with the lead-containing material is avoided until
the treatment of the roasted material in the converters is com-
pleted.

From the data given above, the capacity of each individual
workman is readily determined, as follows: With the reverberatory-
smelting furnace, each man daily works up 0.83 tons; with the
sintering furnace, 1 ton; with the stationary Huntington-Heberlein
furnace, 1.8 tons; with the 6-m. revolving-hearth furnace, 4.5
tons; and with the 8-m. revolving-hearth furnace, 11.8 tons.

A significant change has also taken place in coal consumption.
Thus, when working with the reverberatory and sintering furnaces
in order to attain the requisite temperature of 1000 deg. C.,
there was required not only a comparatively high-grade coal,
but also a large quantity of it. A reverberatory furnace con-
sumed about 503 kg., a sintering furnace about 287 kg., of coal
per ton of ore. For roasting the ore in the stationary and also
in the mechanical Huntington-Heberlein furnaces, a lower tem-
perature (at most 700 deg. C.) is sufficient, as the roasting proper
of the ore is effected in the converters, and the sulphur furnishes
the actual fuel. For this reason, the consumption of coal is
much lower. The comparative figures per ton of ore are as
follows: In the reverberatory furnace, 50.3 per cent.; in the
sintering furnace, 28.7 per cent.; in the stationary Huntington-
Heberlein furnace, 10.3 per cent. ; and in the Huntington-Heberlein
revolving-hearth furnace, 7.3 per cent.

But there is another technical advantage of the Huntington-
Heberlein process which should be mentioned. It is well known
that the volatilization of lead at high temperatures is an exceed-
ingly troublesome factor in the running of a lead-smelting plant;
the recovery of the valuable fume is difficult, and requires con-
densing apparatus, to say nothing of the unhealthful character
of the volatile lead compounds. This volatilization is of course
particularly marked at the high temperatures employed when
working with reverberatory-smelting furnaces; the same is true,
in a somewhat less degree, of the sintering furnaces. In conse-

152 LEAD SMELTING AND REFINING

quence of the markedly lower temperature to which the charge
is heated in the Huntington-Heberlein furnace, and also of the
peculiar mode of completing the roast in blast-converters, the
production of fume is so reduced that the difference between
the values recovered in the old and the new processes is very strik-
ing. Whereas, in 1900, in working up 12,922 tons of ore in the
reverberatory-smelting furnace, and 14,497 tons in the sintering
furnace (27,419 tons in all), there was recovered 2470 tons (or
9 per cent.) as fume from the condensers and smoke flues, the
quantity of fume recovered, in 1903, fell to 879 tons (or 1.8 per
cent.), out of the 48,208 tons of ore roasted, and this notwith-
standing the fact that in the meantime fume-condensing appli-
ances had been considerably expanded and improved, whereby
the collection was much more efficient.

Lastly, the zinc content of the ores no longer exerts the same
unfavorable influence as in the old process (wherein it was advis-
able to subject ore containing much blende to a final washing
before proceeding to the actual metallurgical treatment). In
the new process, the ores are simply roasted without regard to
their zinc content. In this connection it has been found that a
considerable proportion of the zinc passes off with the fume, and
that the roasted material usually contains a quantity of zinc so
small that it no longer causes any trouble in the shaft furnace.
It may also be mentioned here that the ore-dressing plants recently
installed in the mines of Upper Silesia have resulted in a more
perfect separation of the blende.

Shaft Furnaces. — The finished product from the Huntington-
Heberlein blast-converters is of a porous character, and already
contains a part of the flux materials (such as limestone, silica and
iron) which are required for the shaft-furnace charge. It is just
these two characteristics of the roasted product (its porous nature,
on the one hand, leading to its more perfect reduction by the
furnace gases; and, on the other hand, the admixture of fluxes in
the molten condition, resulting in a more complete utilization of
the temperature), which, together with its higher lead and lower
zinc content, determine its ready fusibility. If we further con-
sider that it is possible in the new process to make the total
charge of the shaft furnace richer in lead than formerly (two-
thirds of the total charge as against one-third), and that a higher
blast pressure can be used without danger, it follows immediately

LIME-ROASTING OF GALENA

153

that the capacity of a shaft furnace is much greater by the new
process than by the old method of working. The daily production
of the shaft furnaces on the old and the new process is as shown
in the table given herewith:

TYPE OF SHAFT
FURNACE

CHARACTER OF CHARGE

CHARGE PER DAY,
TONS

WORK-LEAD PRO-
DUCED PER DAY,
TONS

3 tuyeres

("Gray slag from rever-1
| beratory furnaces and |
[ sintered concentrate J

6 to 7

8 tuyeres

3 tuyeres
8 tuyeres

(( tt

[Roasted product of Hunt-1
{ ington-Heberlein process j

(f (C

36 to 38

36
65 to 72

6 to 8

11 to 12
24 to 26

15 tuyeres

(( fl

270

90 to 100

It should be noted that the figure given for the furnace with
15 tuyeres represents the average for 1904; this average is lowered
by the circumstance that during this period there was frequently
a deficiency of roasted material, and the furnace had to work
with low-pressure blast. A truer impression can be gained from
the month of March, 1905, for instance, during which time this
furnace worked under normal conditions; the results are as
follows :

The average for March, 1905, was: Ore charged, 8,269.715
tons; coke, 652.441 tons; total, 8,922.156 tons. Or, in 24 hours:
Ore charged, 266.765 tons; coke, 21.046 tons; total, 287.811 tons.
The production of work-lead was 3,133.245 tons, or 101.069 tons
per day.

The maximum production of roasted ore was 210 tons, on
June 30, 1905, when the total charge was: Ore, 327.38 tons;
coke, 25.2 tons; total, 352.58 tons. The quantity of work-lead
produced on that day was 120.695 tons, while the largest quantity

154 LEAD SMELTING AND REFINING

previously produced in one day was 124.86 tons. It should also
be mentioned that the lead tenor of the slag is almost invariably
below 1 per cent.; it usually lies between 0.3 and 0.5 per cent.

As in the case of the roasting furnaces, the productive capacity
of the shaft furnace also comes out clearly if we figure the number
of furnaces required, on the basis of an annual consumption of
50,000 tons of ore. If we consider 1 ton of the roasted material
as equivalent to 1 ton of ore (which is about right in the case of
the Huntington-Heberlein material, but is rather a high estimate
in the case of the product of the sintering furnace), then, in the
old process (where one-third of the charge was lead-bearing
material), 12 tons could be smelted daily. There would therefore
be needed at least:

50,000 -f- (12 X 300) = 14 three-tuyere shaft furnaces.

Since, as already mentioned, the lead-bearing part of the
charge constitutes two-thirds of the whole in the Huntington-
Heberlein process, the number of shaft furnaces of different types,
as compared with the foregoing, would figure out:

3-tuyere shaft furnace, with product of sintering furnace, 50,000 •*• (12 X

300) = 14 furnaces;
3-tuyere shaft furnace, with product of Huntington-Heberlein furnace,

50,000 -^ (24 X 300) = 7 furnaces;
8-tuyere shaft furnace, with product of Huntington-Heberlein furnace^

50,000 ^ (48 X 300) = 3.4 (say 4) furnaces;

15-tuyere shaft furnace, with product of Huntington-Heberlein furnace,.
50,000 + (180 X 300) =1 furnace.

Running regularly and without interruption, the large shaft
furnace is therefore fully capable of coping with the Huntington-
Heberlein roasted material at the present rate of production.

As regards the number of workmen and the product turned
out per man, no such marked difference is produced by the intro-
duction of the Huntington-Heberlein process in the case of the
shaft furnace as there was noted for the roasting operation.
This is chiefly due to the fact that the work which requires the
more power (such as charging of the furnaces, conveying away
the slag and pouring the lead) can be executed only in part by
mechanical means. Nevertheless, it will be seen from the table
given herewith that, on the one hand, the number of men required

LIME-ROASTING OF GALENA

155

for the charge worked up is smaller; and, on the other, the product
turned out per man has risen somewhat.

«.n

« V.

« M

TYPE OF

s§

W H

a§

gstn

SHAFT

CHARACTER OF CHARGE

§H

o « §

FURNACE

« .j

3 tuyere

Sintered concentrate and gray slag

from reverberatory furnace.

6.0

1.0

8 tuyere
3 tuyere

Gray slag from reverberatory furnace.
Huntington-Heberlein product.

38
36

6
6

6.3
6.0

8
12

1.3

2.0

8 tuyere

Huntington-Heberlein product.

6.0

2.1

15 tuyere

Huntington-Heberlein product.

270

7.9

2.6

A slight difference only is produced by the new process in the
consumption of coke; the economy is a little over 1 per cent.,
the coke consumed being reduced from 9.39 per cent, to 8.17 per
cent, of the total charge. But with the high price of coke, even
this small difference represents a considerable lowering of the
cost of production.

With the great increase in the blast pressure, it would be
supposed that the losses in fume would be much greater than
with the former method of working. But this is not the case;
on the contrary, all experience so far shows that there is much
less fume developed. In 1904, for instance, the shaft-furnace
fume recovered in the condensing system amounted to only
1.06 per cent, of the roasted material, or 0.64 per cent, of the
total charge, as against 2.03 and 1.0 per cent., respectively, in
former years. The observations made on the quantity of flue
dust carried away with the gases escaping into the air through
the stack showed that it is almost nil.

Now, from the loss in fume being slight, from the tenor of
lead in the slag being low, and, on the one hand, from the quantity
of lead-matte produced being much less than before, while on
the other the losses in roasting the ore are greatly reduced —
from all these considerations, it is clear that the total yield must
have been much improved. As a matter of fact, the yield of
lead and silver has been increased by at least 6 to 8 per cent.

Economic Results. — As regards the economical value of the
new process, for obvious reasons no data can be furnished of the
exact expenditure, i.e., the actual total cost of roasting and

156 LEAD SMELTING AND REFINING

smelting the ore. But this at least is placed beyond doubt by
what has been developed above, namely, that considerable saving
must be effected in the roasting, and especially in the smelting,
as compared with the former mode of working. If we take into
account only the economy which is gained in wages through the
increase in the material which one workman can handle, and that
resulting from the reduced consumption of coal and coke, these
alone will show sufficiently that an important diminution of
working cost has taken place. The objection which might be
raised, that the saving effected by reducing manual labor may
be neutralized by the expense of mechanical power (actuating
the roasters, furnishing the compressed blast, etc.), cannot be
regarded as justified, as the cost of mechanical work is com-
paratively low. Thus, for instance, the large 8-m. furnace and
the small, round furnaces require 15 h.p. if worked by electricity,
According to an exact calculation, the cost (to the producer) of
the h.p.-hour, inclusive of machinery, figures out to 3.6 pfennigs
(0.9c.); hence the daily expense for running the revolving-hearth
furnaces amounts to: 15 X 3.6 pfg. X 24 = 12.96 marks ($3.42).
As the seven furnaces together work up: (6 X 27) + 55 = 217
tons of ore, the cost per ton of ore is about 0.06 mark (1.5c.).

The requisite blast is produced by means of single-compression
Encke blowers, of which one is quite sufficient when running at
full load, and then consumes 34 h.p. The daily expenses are
accordingly: 34 X 3.6 pfg. X 24 = 29.28 marks ($7.32); or per
ton of ore, 29.28 -v- 217 = 0.14 mark (3.5c.). Therefore the total
expense for the mechanical work in roasting the ore amounts to
0.06 + 0.14 = 0.20 mark (5c.).

However, the cost of roasting is much more affected by the
expense for keeping the furnaces in repair; another important
factor is the acquisition and maintenance of the tools. Both in
the case of the sintering and also the reverberatory-smelting
furnace, the cost of keeping in repair was high; the consumption
of iron was especially large, owing to the rapid wear of the tools.
This was not surprising, considering that a notably higher tem-
perature prevailed in the reverberatory and sintering furnaces
than in the new roasters, in which the temperature strictly ought
not to rise above 700 deg. C. But in the old type of furnace the
high temperature and the constant working with the iron tools
caused their rapid wear, thus creating a large item for iron and

LIME-ROASTING OF GALENA

157

steel and smith work. In the new process (and more especially
in the revolving-hearth roasters) this disadvantage does not arise.
In this case there is practically no work on the furnace, and the
wear and tear of iron is small. Also, the cost of keeping the
furnaces in repair when working regularly is small as compared
with the old process. In the year 1900, for instance, the cost of
maintenance and tools for the reverberatory and sintering fur-
naces came to 20,701.93 marks ($5,175.48) for treating 27,419.75
tons of ore. Per ton of ore, this represents 0.75 mark (19c.). In
the year 1903, on the other hand, only 9,074.17 marks ($2,268.54)
were expended, although 48,208 tons of ore were worked up in
the three stationary and six mechanical Huntington-Heberlein
furnaces. The cost of maintenance was, therefore, in this case
0.18 mark (4.5c.) per ton of ore.

In the cost of smelting in the shaft furnace, only a slight
difference in favor of the Huntington-Heberlein process is found
if the estimate is based on the total charge; but a marked difference
is shown if it is referred to the lead-bearing portion of the charge,
or to the work-lead produced. Thus the cost of maintenance and
total cost of smelting, figured for one ton of ore, without taking
into account general expenses, have been tabulated as follows:

REDUCTION i

N EXPENSES

PER TON OF

TOTAL
CHARGE

LEAD ORE

WORK-LEAD

(o) Cost of inaiinteDa.DC6 . ...

001M

0.38M

0.67M

(6) Total cost of smelting . . ....

(0.25c)
020M

(9.5c)
6.46M

(16.75c)
11.48M

(5c)

($1.615)

($2.87)

The marked reduction in the expenses, as referred to the
lead-ore and the work-lead produced, is determined (as was
pointed out above) by the greater lead content of the charge ,,
and by the larger yield of lead consequent thereon. The advan-
tage of longer smelting campaigns (which ultimately were mostly
prolonged to one year) also makes itself felt; it would be still
more marked, if the shaft furnace (which was still in working
condition after it was blown out) had been run on for some time
longer.

Finally, if we examine the question of the space taken up by

158 LEAD SMELTING AND REFINING

the plant (which, owing to the scarcity of suitably located building
sites, would have been important at the Friedrichshutte at the
time when the quantity of ore treated was suddenly doubled),
here again we shall recognize the great advantage which this
establishment has gained from the Huntington-Heberlein process.

As was calculated above, there would have been required
15 reverberatory and 15 sintering furnaces to cope with the
quantity of ore treated. As a reverberatory requires, in round
numbers, 120 sq. m. (1290 sq. ft.), and a sintering furnace 200
sq. m. (2153 sq. ft.); and as fully 100 sq. m. (1080 sq. ft.) must
be allowed for each furnace for a dumping ground, therefore the
15 reverberatory furnaces would have required an area of 15 X
120 + 15 X 100 = 3300 sq. m.; the 15 sintering furnaces would
have required 15 X 200 + 15 X 100 = 4500 sq. m.; in all 3300
+ 4500 = 7800 sq. m. (83,960 sq. ft.). The 12 stationary
Huntington-Heberlein furnaces (built together two and two)
would take up a space of 6 X 200 + 12 X 100 = 2400 sq. m.
(25,830 sq. ft.). Similarly, 8 small furnaces would require
8 X 100 + 8 X 100 = 1600 sq. m. (17,222 sq. ft.); while for the
new installation of four 8-meter revolving-hearth furnaces and
10 large converters, only 1320 sq. m. (14,120 sq. ft.) have been
allowed.

For shaft furnaces with three or eight tuyeres, which were
run with low-pressure blast for the material roasted on the old
plan, the total area built upon was 18 X 16.5 = 297 sq. m.;
while a further area of 18 X 14 = 250 sq. m. was hitherto pro-
vided, and was found sufficient for dumping slag when working
regularly. Therefore, the installation of shaft furnaces formerly
in existence, after requisite enlargement to 14 furnaces, would
have demanded a space of 7 X 297 + 7 X 250 = 3829 sq. m.
(42,215 sq. ft.). If four of the small shaft furnaces had been
reconstructed for eight tuyeres, and run with Huntington-Heber-
lein roasted material, using high-pressure blast, the area occupied
would have been reduced to 2 X 297 + 2 X 250 sq. m. = 1094
sq. m. (11,776 sq. ft.).

Still more favorable are the conditions of area required in
the case of the large shaft furnace. This furnace stands in a
building covering an area of 350 sq. m. (3767 sq. ft.), which is
more than sufficient room. The slag-yard (situated in front of
this building, and amply large enough for 36 hours' run) has an

LIME-ROASTING OF GALENA 159

area of 250 sq. m. (2691 sq. ft.); thus the space occupied by the
large shaft furnace, including a yard of 170 sq. m. (1830 sq. ft.),
is in all 780 sq. m. (8396 sq. ft.).

After completion of the new roasting plant and the large
shaft furnace in connection with it, there would be occupied
1320 + 780 = 2100 sq. m. (2260 sq. ft.); and if the system of
reverberatory and sintering furnaces had been continued (with
the requisite additions thereto and to the old shaft-furnace
system), there would have been required 11,629 sq. m. (125,214
sq. ft.). In the estimate above given no regard has been paid
to any of the auxiliary installations (dust chambers, etc.), which,
just as in the case of the old process, would have had to be provided
on a large scale.

It is of course self-evident that both the principal and the
auxiliary installations in the old process would not only have
involved a high first cost, but would also, on account of their
extensive dimensions, have caused considerably greater annual
expense for maintenance.

THE HUNTINGTON-HEBERLEIN PROCESS FROM THE
HYGIENIC STANDPOINT1

BY A. BlERNBAUM
(October 14, 1905)

With regard to the hygienic improvements which the Hun-
tington-Heberlein process offers, we must first deal with the
questions: What were the sources of danger in the old process,
and in what way are these now diminished or eliminated? The
only danger which enters into consideration is lead-poisoning,
other influences detrimental to health being the same in one
process as the other.

With the reverberatory-smelting and roasting-sintering fur-
naces, the chief danger of lead-poisoning lies in the metallic vapor
evolved during the withdrawal of the roasted charge from the
furnace. It is true that appliances may be provided, by which
these vapors are drawn off or led back into the furnace during
this operation; but, even working with utmost care, it is impossible
to insure the complete elimination of lead fumes, especially in
wheeling away the pots filled with the red-hot sintered product.
Moreover, the work at the reverberatory-smelting and roasting-
sintering furnaces involves great physical exertion, wherefore
the respiratory organs of the workmen are stimulated to full
activity, while the exposure to the intense heat causes the men
to perspire freely. Hence, as has been established medically,
the absorption of the poisonous metallic compounds (which are
partially soluble in the perspiration) into the system is favored
both by inhalation of the lead vapor and by its penetration into
the pores of the skin, opened by the perspiration.

A further danger of lead-poisoning was occasioned by the
frequently recurring work of clearing out the dust flues. The
smoke from the reverberatory-smelting furnace especially con-
tained oxidized lead compounds, which on absorption into the

1 Translated from the Zeitschrift fur das Berg.- Hutten- und Salinenwesen
tm. preuss. Staate, 1905, LIU, ii, pp. 219-230.

160

LIME-ROASTING OF GALENA 161

human body might readily be dissolved by the acids of the
stomach, and thus endanger the health of the workmen.

In the Huntington-Heberlein furnaces, on the other hand,
although the charge is raked forward and turned over by hand,
it is not withdrawn, as in the old furnaces, by an opening situated
next to the fire, but is emptied at a point opposite into the con-
verters which are placed in front of the furnace. Moreover, the
converters are filled with the charge at a much lower temperature.
Inasmuch as this charge has already cooled down considerably,
there can be practically no volatilization of lead. The small
quantity of gas which may nevertheless be evolved is drawn off
by fans through hoods placed above the converters.

A further improvement, from the hygienic point of view, is in
the use of the mechanical furnaces, from which the converters
can be filled automatically (almost without manual labor, and
with absolute exclusion of smoke). The converters are then
placed on their stands and blown. This work also is carried out
under hoods, as gas-tight as possible, furnished with a few closable
working apertures. During the blowing of the material, the
work of the attendant consists solely in keeping up the charge
by adding more cold material and filling any holes that may be
formed. It does not entail nearly as much physical strain as
the handling of the heavy iron tools and the continued exposure
of the workmen to the hottest part of the furnace, which the
former roasting process involved.

Some experiments carried out with larger converters (of 4-
and 10-ton capacity) have indicated the direction in which the
advantages mentioned above may probably be developed to such
a point that the danger of lead-poisoning need hardly enter into
consideration. Both the charging of the revolving-hearth fur-
naces and the filling of the converters are to be effected mechani-
cally. Furthermore, in the case of the large converters the
filling up of holes becomes unnecessary, and no manual work of
any kind is required during the whole time of blowing. The
converters can be so perfectly enclosed in hoods that the escape
of gases into the working-rooms becomes impossible, and lead-
poisoning of the men can occur only under quite unusual circum-
stances.

The beneficial influence on the health of the workmen attend-
ing on the roasting furnaces, occasioned by the introduction of

162

LEAD SMELTING AND REFINING

the Huntington-Heberlein process, can be seen from the statistics
of sickness from lead-poisoning for the years 1902 to 1904, as
given herewith:

LEAD -POISONING

CASES CON-
TRACTED

"n c

CASES

NESS

_•

METHOD OF WORKING

YEAR

O "3

0 2

££

fjj

Old

( 1902
I 1903

15
12

16.1
13.9

246
222

264.5
258.1

4
5

H -H

1904

9.2

242

278.2

This shows a gratifying decrease in the number of cases,
namely, from 16.1 to 9.2 per cent.; this decrease would have been
still greater if Huntington-Heberlein furnaces had been in use
exclusively. However, most of the time two or three sintering
furnaces were fired for working up by-products, 16 to 18 men
being engaged on that work. The Huntington-Heberlein furnaces
alone (at which, in the year 1904, 69 men in all were occupied)
show only 2.9 per cent, of cases. That the number of days of
illness was not reduced is due to the fact that the cases among
the gang of men working at the sintering furnaces were mostly
of long standing and took some time to cure.

The noxious effects upon the health of the workmen in running
the shaft furnaces are due to the fumes from the products made
in this operation, such as work-lead, matte and slag, which flow
out of the furnace at a temperature far above their melting points.
Even with the old method of running the shaft furnaces the
endeavor has always been to provide as efficiently as possible
against the danger caused by this volatilization, and, wherever
feasible, to install safety appliances to prevent the escape of lead
vapors into the work-rooms; but these measures could not be
made as thorough as in the case of the Huntington-Heberlein
process.

The principal work in running the shaft furnaces, aside from
the charging, consists in tapping the slag and pouring out the
work-lead. Other unpleasant jobs are the barring down (which

LIME-ROASTING OF GALENA 163

In the old process had to be done frequently) and the cleaning
out of the furnace after blowing out.

In the old process the slag formed in the furnace flows out
continuously through the tap-hole into iron pots placed in front
of the spout. A number of such pots are so arranged on a revol-
ving table that as soon as one is filled the next empty can be brought
up to the duct; thus the slag first poured in has time to cease
fuming and to solidify before it is removed. The vapors arising
from the slag as it flows out are conveyed away through hoods.
At the same time with the slag, lead matte also issues from the
furnace. Now the greater the quantity of lead matte, the more
smoke is also produced; and, with the comparatively high pro-
portion of lead matte resulting from the old process, the quantity
of smoke was so great that the ventilation appliances were no
longer sufficient to cope with it, thus allowing vapors to escape
into the work-room.

The work-lead collects at the back of the furnace in a well,
from which it is from time to time ladled into molds placed near
by. If the lead is allowed to cool sufficiently in the well, it does
not fume much in the ladling out. But when the furnace runs
very hot (which sometimes happens), the lead also is hotter and
is more inclined to volatilize. In this event the danger of lead-
poisoning is very great, for the workman has to stand near the
lead sump.

A still greater danger attends the work of barring down and
cleaning out the furnace. The barring down serves the purpose
of loosening the charge in the zone of fusion; at the same time
it removes any crusts formed on the sides of the furnace, or
obstructions stopping up the tuyeres. With the old furnaces,
and their strong tendency to crust, this work had to be under-
taken almost every day, the men being compelled to work for
rather a long time and often very laboriously with the heavy iron
tools in the immediate neighborhood of the glowing charge, the
front of the furnace being torn open for this purpose. In this
operation they were exposed without protection to the metallic
.vapors issuing from the furnace, inasmuch as the ventilating
appliances had to be partially removed during this time, in order
to render it at all possible to do the work.

In a similar manner, but only at the time of shutting down
& shaft furnace, the cleaning out (that is to say, the withdrawing

164 LEAD SMELTING AND REFINING

of no longer fused but still red-hot portions of the charge left in
the furnace) is carried out. In this process, however, the glowing
material brought out could be quenched with cold water to such
a point that the evolution of metallic vapors could be largely
avoided.

Lastly, the mode of charging of the shaft furnace is also to be
regarded as a cause of poisoning, inasmuch as it is impossible to
avoid entirely the raising of dust in the repeated act of dumping
and turning over the materials for smelting, in preparing the mix,
and in subsequently charging the furnace.

By the introduction of the Huntington-Heberlein process, all
these disadvantages, both in the roasting operation and in running
the shaft furnaces, are in part removed altogether, in part reduced
to such a degree that the danger of injury is brought to a minimum.

In furnaces in which the product of the Huntington-Heberlein
roast is smelted, the slag is tapped only periodically at considerable
intervals; and, as there is less lead matte produced than formerly,
the quantity of smoke is never so great that the ventilating fan
cannot easily take care of it. There is therefore little chance of
any smoke escaping into the working-room.

As the production of work-lead, especially in the case of the
large shaft furnace, is very considerable, so that the lead contin-
ually flows out in a big stream into the well, the hand ladling has
to be abandoned. Therefore the lead is conducted to a large
reservoir standing near the sump, and is there allowed to cool
below its volatilizing temperature. As soon as this tank is full,
the lead is tapped off and (by the aid of a swinging gutter) is cast
into molds ready for this purpose. Both the sump and the
reservoir-tank are placed under a fume-hood. The swinging
gutter is covered with sheet-iron lids while tapping, so that any
lead volatilized is conveyed by the gutter itself to a hood attached
to the reservoir; thus the escape of metallic vapors into the
working space is avoided, as far as possible.

This method of pouring does not entail the same bodily exer-
tion as the ladling of the lead; moreover, as it requires but little
time, it gives the workmen frequent opportunity to rest.

But one of the chief advantages of the Huntington-Heberlein
process lies in the entire omission of the barring down. If the
running of the shaft furnace is conducted with any degree of care,
disorders in the working of the furnace do not occur, and one

LIME-ROASTING OF GALENA

165

can rely on a perfectly regular course of the smelting process
day after day. No formation of any crusts interfering with the
operation of the furnace has been recorded during any of the
campaigns, which have, in each case, lasted nearly a year.

As regards the cleaning out of the furnace, this cannot be
avoided on blowing out the Huntington-Heberlein shaft furnace;
but at most it occurs only once a year, and can be done with less
danger to the workmen, owing to the better equipment.

Further, the charge is thrown straight into the furnace (in the
case of the large shaft furnace); thus the repeated turning over
of the smelting material, as formerly practised, becomes unneces-
sary, and the deleterious influence of the unavoidable formation
of dust is much diminished.

The accompanying statistics of sickness due to lead-poisoning
in connection with the operation of the shaft furnace (referring
to the same period of time as those given above for the roasting
furnaces) confirm the above statements.

YEAR

No. OF MEN

LEAD-POISONING — SHAFT FURNACES

CASES

DAYS OF ILLNESS

TOTAL

PER 100 PER-
SONS

TOTAL

PER 100 PER-
SONS

1902
1903
1904

250
267
232

58
59
24

23.2
22.1
10.3

956
1044
530

382.4
391.0

228.4

If it were possible to make the necessary distinctions in the
case of the large shaft furnace, the diminution in sickness from
lead-poisoning would be still more apparent; for, among the fur-
nace attendants proper, there has been no illness; all cases of
poisoning have occurred among the men who prepare the charge
who break up the roasted material, and others who are occupied
with subsidiary work. Some of these are exposed to illness
through their own fault, owing to want of cleanliness, or to
neglect of every precautionary measure against lead-poisoning.

Thus far we have dealt only with the advantages and improve-
ments of the Huntington-Heberlein process; we will now, in
conclusion, consider also its disadvantages.

The chief drawback of the new process lies in the difficulty of
breaking up the blocks of the roasted product from the con-

166 LEAD SMELTING AND REFINING

yerters, a labor which, apart from the great expense involved, is
also unhealthy for the workmen engaged thereon. Seemingly this
evil is still further increased by working with larger charges in
the 10-ton converters, as projected; but in this case it is proposed
to place the converters in an elevated position, and to cause the
blocks to be shattered by their fall from a certain hight, so that
further breaking up will require but little work. Trials made in
this direction have already yielded satisfactory results, and seem
to promise that the disadvantage will in time become less im-
portant.

Another unpleasant feature is the presence (in the waste gases
from the converters) of a higher percentage of sulphur dioxide,
the suppression of which, if it is feasible at all, might be fraught
with trouble and expense.

That the roaster gases from the reverberatory-smelting and
sintering furnaces did not show such a high percentage of sulphur
dioxide must be ascribed chiefly to the circumstance that the
roasting was much slower, and that the gases were largely diluted
with air already at the point where they are formed, as the work
must always be done with the working-doors open. In the
Huntington-Heberlein process, on the other hand, the aim is to
prevent, as far as possible, the access of air from outside while
blowing the charge. The more perfectly this is effected, and the
greater the quantity of ore to be blown in the converters, the
higher will also be the percentage of sulphur dioxide in the waste
gases. This circumstance has not only furnished the inducement,
but it has rendered it possible to approach the plan of utilizing
the sulphur dioxide for the manufacture of sulphuric acid. If
this should be done successfully (which, according to the experi-
ments carried out, there is reasonable ground to expect), the
present disadvantage might be turned into an advantage. This
has the more significance because an essential constituent of the
lead ore — the sulphur — will then no longer, as hitherto, have
to be regarded as wholly lost.1

lThe manufacture of sulphuric acid from these gases has now been
undertaken in Silesia on a working scale. — EDITOR.

THE HUNTINGTON-HEBERLEIN PROCESS

BY THOMAS HUNTINGTON AND FERDINAND HEBERLEIN

(May 26, 1906)

This process for roasting lead sulphide ores has now fairly
established itself in all parts of the world, and is recognized by
metallurgical engineers as a successful new departure in the
method of desulphurization. It offers the great advantage over
previous methods of being a more scientific application of the
roasting reactions (of the old well-used formulae PbS +. 3O =
PbO 4- SO2 and PbS + PbSO4 + 2O = 2PbO + 2SO2) and ad-
mits of larger quantities being handled at a time, so that the use
of fuel and labor are in proportion to the results achieved, and
also there is less waste all around in so far as the factors necessary
for the operation — fuel, labor and air — can be more economi-
cally used. The workman's time and strength are not employed
in laboriously shifting the ore from one part of the furnace to
another with a maximum amount of exertion and a minimum
amount of oxidation. The fuel consumed acts more directly
upon the ore during the first part of the process in the furnace
and its place is taken by the sulphur itself during the final and
blowing stage, so that during the whole series of operations more
concentrated gases are produced and consequently the large excess
of heated air of the old processes is avoided to such an extent that
the gases can be used for the production of sulphuric acid.

With a modern well-constructed plant practically all the evils
of the old hand-roasting furnaces are avoided, and besides the
notable economy achieved by the H.-H. process itself, the health
and well-being of the workmen employed are greatly advanced,
so that where hygienic statistics are kept it is proved that lead-
poisoning has greatly diminished. It is only natural, therefore,
that the H.-H. process should have been a success from the start,
popular alike with managers and workmen once the difficulties
inseparable from the introduction of any new process were over-
come.

167

168 LEAD SMELTING AND REFINING

Simple as the process now appears, however, it is the result
of many years of study and experiment, not devoid of disap-
pointments and at times appearing to present a problem incapable
of solution. The first trials were made in the smelting works at
Pertusola, Italy, as far back as 1889, where considerable sums
were devoted every year to this experimental work and lead ore
roasting was almost continuously on the list of new work from
1875 on.

It may be interesting to mention that at this time the Monte-
vecchio ores (containing about 70 per cent, lead and about 15
per cent, sulphur, together with a certain amount of zinc and
iron) were considered highly refractory to roast, and the only
ores approved of by the management of the works at this date
were the Monteponi and San Giovanni first-class ores (containing
about 80 per cent, lead), and the second-class carbonates (with
at least 60 per cent, lead and 5 per cent, sulphur). It must be
noted that a modified Flintshire reverberatory process was in use
in the works, which could deal satisfactorily only with this class
of ore, so that, as these easy ores diminished in quantity every
year and their place was taken by the " refractory" Montevecchio
type, the roasting problem was always well to the front at the
Pertusola works.

It may be asserted that almost every known method of desul-
phurization was examined and experimented upon on a large
scale. Gas firing was exclusively used on certain classes of ores
for several years with considerable success, and revolving furnaces
of the Bruckner type — gas fired — were also tried. Although
varying degrees of success were obtained, no really great progress
was made in actual desulphurization; methods were cheapened
and larger quantities handled at a time, but the final product —
whether sintered or in a pulverulent state — seldom averaged
much under 5 per cent, sulphur, while the days of the old "gray
slags" (1 per cent, to 2 per cent, sulphur) from the reverberatories
totally disappeared, together with the class of ores which produced
them.

During the long period of these experiments in desulphurization
various facts were established:

(1) That sulphide of lead — especially in a pulverulent state
— could not be desulphurized in the same way as other sulphides,
such as sulphides of iron, copper, zinc, etc., because if roasted in

LIME-ROASTING OF GALENA 169

a mechanical furnace the temperature had to be kept low enough
to avoid premature sintering, which would choke the stirrers and
cause trouble by the ore clogging on the sides and bottom of the
furnace. If, however, the ore was roasted in a "dry state" at
low temperature, a great deal of sulphur remained in the product
as sulphate of lead, which was as bad for the subsequent blast-
furnace work as the sulphide of lead itself. When air was pressed
through molten galena — in the same way as through molten
copper matte — a very heavy volatilization of lead took place,
while portions of it were reduced to metal or were contained as
sulphide in the molten matte, so that a good product was not
obtained.

(2) That no complete dead roast of lead ores could be obtained
unless the final product was thoroughly smelted and agglomerated.

(3) That a well roasted lead ore could be obtained by oxidizing
the PbS with compressed air, after the ore had been suitably
prepared.

(4) That metal losses were mainly due to the excessive heat
produced in the oxidation of PbS to PbO, and other sulphides
present in the ore.

It was by making use of these facts that the H.-H. roasting
process was finally evolved, and by carefully applying its principles
it is possible to desulphurize completely the ore to a practically
dead roast of under 1 per cent, sulphur; in practice, however,
such perfection is unnecessary and a well agglomerated product
with from 2 to 2.5 per cent, sulphur is all that is required. During
some trials in Australia, where a great degree of perfection was
aimed at, a block of over 2000 tons of agglomerated, roasted ore
was produced containing 1 per cent, sulphur (as sulphide); as
the ores contained an average of about 10 per cent. Zn, this was
a very fine result from a desulphurization point of view, but it
was not found that this 1 per cent, product gave any better results
in the subsequent smelting in the blast furnace than later on a
less carefully prepared material containing 2.5 per cent, sulphur.

In the early stages of experiment the great difficulty was to
obtain agglomeration without first fusing the sulphides in the
ore, and turning out a half-roasted product full of leady matte.
Simple as the thing now is, it seemed at times impossible to avoid
this defect, and it was only by a careful study of the effects of an
addition of lime, Fe2O3 or Mn2O3, and their properties that the

170 LEAD SMELTING AND REFINING

right path was struck. Before the introduction of the H.-H. pro-
cess lime was only used in the reverberatory process (Flintshire and
Tarnowitz) to stiffen the charge, but as Percy tells us that after
its addition the charge was glowing, it must have had a chemical
as well as a mechanical effect. In recognition of this fact fine
caustic lime or crushed limestone was mixed with the ore before
charging it into the furnace and exposing it to an oxidizing heat.

It was thought probable that a dioxide of lime might be tem-
porarily formed, which in contact with PbS would be decomposed
immediately after its formation, or that the CaO served as Con-
tactsubstanz in the same way as spongy platinum, metallic silver,
or oxide of iron. As CaSO4 and not CaSO3 is always found in
the roasted ore, this may prove that CaO is really a contact
substance for oxygen (see W. M. Hutchings, Engineering and
Mining Journal, Oct. 21, 1905, Vol. LXXX, p. 726). The fact
that the process works equally well with Fe2O3 instead of CaO
speaks against the theory of plumbate of lime. Whatever theory
may be correct, the fact remains that CaO assists the roasting pro-
cess and that by its use the premature agglomeration of the sul-
phide ore is avoided. A further advantage of lime is that it keeps
the charge more porous and thus facilitates the passage of the air.

The shape and size of the blowing apparatus best adapted for
the purpose in view occupied many months; starting from very
shallow pans or rectangular boxes several feet square with a few
inches of material over a perforated plate, it gradually resolved
•itself into the cone-shaped receptacle — holding about a ton of
ore — as first introduced together with the process. In later
years and in treating larger quantities a more hemispherical form
has been adopted, containing up to 15 tons of ore.

It is probable about eight years were employed in actually
working out the process before it was introduced on any large
scale at Pertusola, but by the end of 1898 the greater part of the
Pertusola ores were treated by the process. Its first introduction
to any other works was in 1900, when it was started outside its
home for the first time at Braubach (Germany). Since then
its application has gradually extended, proceeding from Europe
to Australia and Mexico and finally to America and Canada,
where recognition of its merits was more tardy than elsewhere.
It is now practically in general use all over the world and is
recognized as a sound addition to metallurgical progress. It is

LIME-ROASTING OF GALENA 171

doubtless only a step in the right direction and with its general
use a better knowledge of its principles will prevail, so that its
future development in one direction or another, as compared
with present results, may show some further progress.

The present working of the H.-H. process still follows prac-
tically the original lines laid down, and by preliminary roasting
in a furnace with lime, oxide of iron, or manganese (if not already
contained in the ore), prepares the ore for blowing in the converter.
Mechanical furnaces have been introduced to the entire exclusion
of the old hand-roasters, and the size of the converters has been
gradually increased from the original one-ton apparatus succes-
sively to 5-, 7- and 10-ton converters; at present some for 15 tons
are being built in Germany and will doubtless lead to a further
economy.

The mechanical furnace at present most in use is a single-
hearth revolving furnace with fixed rabbles, the latest being
built with a diameter of 26J ft. and a relatively high arch to
ensure a clear flame and rapid oxidation of the ore. The capacity
of these furnaces varies, of course, with the nature of the ores to
be treated, but with ordinary lead ores (European and Australian
practice) of from 50 per cent, to 60 per cent, lead and 14 per
cent, to 18 per cent, sulphur, the average capacity may be taken
at about 50 to 60 tons of crude ore per day of 24 hours. The
consumption of coal with a well-constructed furnace is very low
and is always under 8 per cent. — 6 per cent, being perhaps the
average. These furnaces require very little attention, being
automatic in their charging and discharging arrangements.

The ore on leaving the furnace is charged into the converters
by various mechanical means (Jacob's ladders, conveyors, etc.).
The converter charge usually consists of some hot ore direct
from the furnace, on top of which ore is placed which has been
cooled down by storage in bins or by the addition of water. The
converter is generally filled in two charges of five tons each, and
the blowing time should not be more than 4 to 6 hours. The
product obtained should be porous and well agglomerated, but
easily broken up, tough melted material being due to an excess
of silica and too much lead sulphide. Attention, therefore, to
these two points (good preliminary roasting and correction of the
charge by lime) obviates this trouble. This roasted ore should not
contain more than about 1.5 to 2 per cent, sulphur, and in a

172 LEAD SMELTING AND REFINING

modern blast furnace gives surprisingly good results, the matte-
fall being in most cases reduced to nothing, and the capacity of
the furnace is largely increased, while the slags are poorer.

If the converter charge has been properly prepared, the blow-
ing operation proceeds with the greatest smoothness and requires
very little attention on the part of the workmen, the heat and
oxidation rise gradually from the bottom and volatilization losses
remain low, so that it is possible, if desired, to produce hot con-
centrated sulphurous gases suitable for the manufacture of
sulphuric acid.

Besides the actual economy obtained in roasting ores by the
process, a great feature of its success has been the remarkable
improvement in smelting and reducing the roasted ore as com-
pared with previous experience. This is due to the nature of
the roasted material, which, besides being much poorer in sulphur
than was formerly the case, is thoroughly porous and well ag-
glomerated and contains — if the original mixture is properly
made — all the necessary slagging materials itself, so that it
practically becomes a case of smelting slags instead of ore, and
to an expert the difference is evident.

Experience has shown that on an average the improvement
in the capacity of the blast furnace may be taken at about 50 to
100 per cent., so that in works using the H.-H. process — after
its complete introduction — about half the blast furnaces for-
merly necessary for the same tonnage were blown out. The
matte-fall has become a thing of the past, so that, except in
those cases where some matte is required to collect the copper
contained in the ores, lead matte has disappeared and the quantity
of flue dust as well as the lead and silver losses have been greatly
reduced.

Referring to the latest history of the H.-H. process, and the
theory of direct blowing, it may be remarked — putting aside all
legal questions — that the idea, metallurgically speaking, is
attractive, as it would seem that by eliminating one-half of the
process and blowing the ores direct without the expense of a
preliminary roast a considerable economy should be effected.
Upon examination, however, this supposed economy and sim-
plicity is not at all of such great importance, and in many cases,
without doubt, would be retrogressive in lead ore smelting rather
than progressive. When costs of roasting in a furnace are reduced

LIME-ROASTING OF GALENA 173

to such a low figure as can be obtained by using 50-ton furnaces
and 10- to 15-ton converters, there is very little margin for im-
provement in this direction. From the published accounts of
the Tarnowitz smelting works (the Engineering and Mining
Journal, Sept. 23, 1905, Vol. LXXX, p. 535) the cost of mechanical
preliminary roasting cannot exceed 25c. per ton, so that even
assuming direct blowing were as cheap as blowing a properly
prepared material, the total economy would only be the above
figure, viz., 25c.; but this is far from being the case.

Direct blowing of a crude ore is considerably more expensive
than dealing with the H.-H. product, because of necessity the
blowing operation must be carried out slowly and with great
care so as to avoid heavy metal losses, and whereas a pre-roasted
ore can be easily blown in four hours and one man can attend to
two or three 10-ton converters, the direct blowing operation takes
from 12 to 18 hours and requires the continual attention of one
man. In the first case the cost of labor would be: One man at
say $3 for 50 tons (at least), i.e., 6c. per ton, and in the second
case one man at $3 for 10 tons (at the best), i.e., 30c., a difference
in favor of pre-roasting of 24c., so that any possible economy
would disappear. Furthermore, as the danger of blowing upon
crude sulphides for 12 or 18 hours is greater as regards metal
losses than a quick operation of four hours, it is very likely that
instead of an economy there would be an increase in cost, owing
to a greater volatilization of metals.

These remarks refer to ordinary lead ores with say 50 per
cent, lead and about 14 per cent, sulphur. With ores, however,
such as are generally treated in the United States the advantages
of pre-roasting are still more evident. These ores contain about
10 to 15 per cent, lead, 30 to 40 per cent, sulphur, 20 to 30 per
cent, iron, 10 per cent, zinc, 5 per cent, silica, and lose the
greater part of the pyritic sulphur in the preliminary roasting,
leaving the iron in the form of oxide, which in the subsequent
blowing operation acts in the same way as lime. For this reason
the addition of extra fluxes, such as limestone, gypsum, etc., to
the original ore is not necessary and only a useless expense.

In certain exceptional cases and with ores poor in sulphur,
direct blowing might be applicable, but for the general run of
lead ores no economy can be expected by doing away with the
preliminary roast.

MAKING SULPHURIC ACID AT BROKEN HILL

(August 11, 1904)

The Broken Hill Proprietary Company has entered upon the
manufacture of sulphuric acid on a commercial scale. The acid
is practically a by-product, being made from the gases emanating
from the desulphurization of the ores, concentrates, etc., by the
Carmichael-Bradford process. The acid can be made at a mini-
mum of cost, and thus materially enhances the value of the process
recently introduced for the separation of zinc blende from the
tailings by flotation. The following particulars are taken from
a recently published description of the process: The ores, concen-
trates, slimes, etc., as the case may be, are mixed with gypsum,
the quantity of the latter varying from 15 to 25 per cent. The
mixture is then granulated to the size of marbles and dumped
into a converter. The bottom of the charge is heated from
400 to 500 deg. C. It is then subjected to an induced current
of air, and the auxiliary heat is turned off. The desulphuriza-
tion proceeds very rapidly with the evolution of heat and the
gases containing sulphurous anhydride. The desulphurization
is very thorough, and no losses occur through volatilization.
The sulphur thus rendered available for acid making is rather
more than is contained in the ore, the sulphur in the agglomerated
product being somewhat less than that accounted for by the
sulphur contained in the added gypsum. Thus from one ton
of 14 per cent, suphide ore it is possible to make about 12 cwt.
of chamber acid, fully equaling 7 cwt. of strong acid.

The plant at present in use, which comprises a lead chamber
of 40,000 cu. ft., can turn out 35 tons of chamber acid per week.
This plant is being duplicated, and it has also been decided to
erect a large plant at Port Pirie for use in the manufacture of
superphosphates. It is claimed that the production of sulphuric
acid from ores containing only 14 per cent, of sulphur establishes
a new record.

174

THE CARMICHAEL-BRADFORD PROCESS

BY DONALD CLARK

(November 3, 1904)

Subsequent to the introduction of the Huntington-Heberlein
process in Australia, Messrs. Carmichael and Bradford, two
employees of the Broken Hill Proprietary Company, patented a
process which bears their name. Instead of starting with lime,
or limestone and galena, as in the Huntington-Heberlein process,
they discovered that if sulphate of lime is mixed with galena
and the temperature raised, on blowing a current of air through
the mixture the temperature rises and the mass is desulphurized.
The process would thus appear to be a corollary of the original
one, and the reactions in the converter are identical. Owing to
the success of the acid processes in separating zinc sulphide from
the tailing at Broken Hill, it became necessary to manufacture
sulphuric acid locally in large quantity. The Carmichael-Bradford
process has been started for the purpose of generating the sulphur
dioxide necessary, and is of much interest as showing how gases
rich enough in SO2 may be produced from a mixture containing
only from 13 to 16 per cent, sulphur.

Gypsum is obtained in a friable state within about five miles
from Broken Hill. This is dehydrated, the CaSO, 2H2O being
converted into CaS04 on heating to about 200 deg. C. The
powdered residue is mixed with slime produced in the milling
operations and concentrate in the proportion of slime 3 parts,
concentrate 1 part, and lime sulphate 1 part. The proportions
may vary to some extent, but the sulphur contents run from
13 to 16 or 17 per cent. The average composition of the ingre-
dients is as given in the table on the next page.

These materials are moistened with water and well mixed by
passing them through a pug-mill. The small amount of water
used serves to set the product, the lime sulphate partly becoming
plaster of paris, 2CaSO, H2O. While still moist the mixture is
broken into pieces not exceeding two inches in diameter and

175

176

LEAD SMELTING AND REFINING

spread out on a drying floor, where excess of moisture is evapo-
rated by the conjoint action of sun and wind.

SLIME

CONCEN-
TRATE

CALCIUM
SULPHATE

AVERAGE

Galena

Blende

Pyrite

Ferric oxide. . . ....

2 5

Ferrous oxide ....

Manganous oxide

6.5

Alumina

5 5

Lime

3 5

Silica .

Sulphur trioxide ....

The pots used are small conical cast-iron ones, hung on trun-
nions, and of the same pattern as used in the Huntington-Heber-
lein process. Three of these are set in line, and two are at work
while the third is being filled. These pots have the same form
of conical cover leading to a telescopic tube, and all are connected
to the same horizontal pipe leading to the niter pots. Dampers
are provided in each case. A small amount of coal or fuel is
fed into the pots and ignited by a gentle blast; as soon as a tem-
perature of about 400 to 500 deg. C. is attained the dried mixture
is fed in, until the pot is full; the cover is closed down and the
mass warms up. Water is first driven off, but after a short time
concentrated fumes of sulphur dioxide are evolved. The amount
of this gas may be as much as 14 per cent., but it is usually
kept at about 10 per cent., so as to have enough oxygen for the
conversion of the dioxide to the trioxide. The gases are led over
a couple of niter pots and thence to the usual type of lead chamber
having a capacity of 40,000 cu. ft. Chamber acid alone is made,
since this requires to be diluted for what is known as the salt-
cake process.

The plant has now been in operation for some time and, it is
claimed, with highly successful results. The product tipped out
of, the converter is similar to that obtained in the Huntington-
Heberlein process, and is at once fit for the smelters, the amount
of sulphur left in it being always less than that originally intro-
duced with the gypsum; analysis of the desulphurized material
shows usually from 3 to 4 per cent, sulphur.

THE CARMICHAEL-BRADFORD PROCESS

BY WALTER RENTON INGALLS

(October 28, 1905)

As described in United States patent No. 705,904, issued
July 29, 1902, lead sulphide ore is mixed with 10 to 35 per cent,
of calcium sulphate, the percentage varying according to the
grade of the ore. The mixture is charged into a converter and
gradually heated externally until the lower portion of the charge,
say one-third to one-fourth, is raised to a dull-red heat; or the
reactions may be started by throwing into the empty converter
a shovelful of glowing coal and turning on a blast of air sufficient
to keep the coal burning and then feeding the charge on top of
the coal. This heating effects a reaction whereby the lead sulphide
of the ore is oxidized to sulphate and the calcium sulphate is
reduced to sulphide. The heated mixture being continuously
subjected to the blast of air, the calcium sulphide is re-oxidized
to sulphate and is thus regenerated for further use. This reaction
is exothermic, and sufficient heat is developed to complete the
desulphurization of the charge of ore by the concurrent reactions
between the lead sulphate (produced by the calcium sulphate)
and portions of undecomposed ore, sulphurous anhydride being
thus evolved. The various reactions, which are complicated in
their nature, continue until the temperature of the charge reaches
a maximum, by which time the charge has shrunk considerably
in volume and has a tendency to become pasty. This becomes
more marked as the production of lead oxide increases, and as
the desired point of desulphurization is attained the mixture
fuses; at this stage the calcium sulphide which is produced from
the sulphate cannot readily oxidize, owing to the difficulty of
coming into actual contact with the air in the pasty mass, but,
being subjected to the strong oxidizing effect of the metallic
oxide, it is converted into calcium plumbate, while sulphurous
anhydride is set free. The mass then cools, as the exothermic
reactions cease, and can be readily removed to a blast furnace
for smelting.

177

178 LEAD SMELTING AND REFINING

The reactions above described are as outlined in the original
American patent specification. Irrespective of their accuracy,
the Carmichael-Bradford process is obviously quite similar to
the Huntington-Heberlein, and doubtless owes its origin to the
latter. The difference between them is that in the Huntington-
Heberlein process the ore is first partially roasted with addition
of lime, and is then converted in a special vessel. In the Car-
michael-Bradford process the ore is mixed with gypsum and is
then converted directly. The greatest claim for originality in
the Carmichael-Bradford process may be considered to lie in it as
a method of desulphurizing gypsum, inasmuch as not only is the
sulphur of the ore expelled, but also a part of the sulphur of
the gypsum; and the sulphur is driven off as a gas of sufficiently
high tenor of sulphur dioxide to enable sulphuric acid to be
made from it economically. Up to the present time the Car-
michael-Bradford process has been put into practical use only at
Broken Hill, N. S. W.

The Broken Hill Proprietary Company first conducted a series
of tests in a converter capable of treating a charge of 20 cwt.
These tests were made at the smelting works at Port Pirie. Ex-
haustive experiments made on various classes of ores satisfac-
torily proved the general efficacy of the process. The following
ores were tried in these preliminary experiments, viz.:

First-grade concentrate containing: Pb, 60 per cent.; Zn, 10
per cent.; S, 16 per cent.; Ag, 30 oz.

Second-grade concentrate containing: Pb, 45 per cent.; Zn,
12.5 per cent.; S, 14.5 per cent.; Ag, 22 oz.

Slime containing: Pb, 21 per cent.; Zn, 17 per cent.; S, 13 per
cent.; Ag, 18 oz.

Lead-copper matte containing: Fe, 42 per cent.; Pb, 17 per
cent.; Zn, 13.3 per cent.; Cu, 2.4 per cent.; S, 23 per cent.; Ag;
25 oz.

Other mattes, of varying composition up to 45 per cent. Pb
and 100 oz. Ag, were also tried.

The results from these preliminary tests were so gratifying
that a further set of tests was made on lead-zinc slime, with a
view of ascertaining whether any volatilization losses occurred
during the desulphurization. This particular material was chosen
because of its accumulation in large proportions at the mine,
and the unsatisfactory result of the heap roasting which has

LIME-ROASTING OF GALENA

179

recently been practised. The heap roasting, although affording
a product containing only 7 per cent. S, which is delivered in
lump form and therefore quite suitable for smelting, resulted in
a high loss of metal by volatilization (17 per cent. Pb, 5 per cent.
Ag).

The result of nine charges of the slime treated by the Car-
michael-Bradford process was as follows:

Cwt

Ass

AYS

Cow

fENTS

Pb%

Agoz.

Zn%

Pb
cwt.

Ag.
oz.

Zn
cwt.

S
cwt.

Raw slime

128.1

21.3

18.0

16.8

13.1

27.28

115.3

25.2

16.78

Raw gypsum

54.9

9.88

Total

183.0

27.28

115.3

25.2

26.66

109 88

207

17.2

4.80

22.74

94.5

527

Middling

14.47

17.7

15.7

6.20

2.56

11.3

0.89

Fines.

11 12

190

14.8

7.50

2.11

8.2

0.83

Total

135.47

5.17

27.41

113.0

6.99

These results indicated practically no volatilization of lead
-and silver during the treatment, the lead showing a slight increase,
viz., 0.47 per cent., and the silver 1.13 per cent. loss. A desul-
phurization of 70.4 per cent, was effected. A higher desulphuri-
zation could have been effected had this been desired. In the
above tabulated results, the term " middling" is applied to the
loose fritted lumps lying on the top of the charge: these are suitable
for smelting, the fines being the only portion which has to be
returned.

In order to test the practicability of making sulphuric acid,
a plant consisting of three large converters of capacity of five
tons each, together with a lead chamber 100 ft. by 20 ft. by 20 ft.,
was then erected at Broken Hill, together with a dehydrating
furnace, pug-mill, and granulator. These converters are shown
in the accompanying engravings.

A trial run was made with 108 tons of concentrate of the
following composition: 54 per cent, lead; 1.9 per cent, iron; 0.9
per cent, manganese; 9.4 per cent, zinc; 14.6 per cent, sulphur;
19.2 per cent, insoluble residue, and 24 oz. silver per ton.

The converter charge consisted of 100 parts of the concentrate
and 25 parts of raw gypsum, crushed to pass a 1-in. hole and

180 LEAD SMELTING AND REFINING

retained by a 0.25-in. hole, the material finer than 0.25 in. (which
amounted to 5 per cent, of the total) being returned to the pug-
mill. After desulphurization in the converter, the product as-
sayed as follows: 48.9 per cent, lead; 1.80 per cent, iron; 0.80 per
cent, manganese; 7.87 per cent, zinc; 3.90 per cent, sulphur;
1.02 per cent, alumina; 5.80 per cent, lime; 21.75 per cent, insoluble
residue; 8.16 per cent, undetermined (oxygen as oxides, sulphates,
etc.); total, 100 per cent. Its silver content was 22 oz. per ton.
The desulphurized ore weighed 10 per cent, more than the raw
concentrate. During this run 34 tons of acid were made.

A trial was then made on 75 tons of slime of the following
composition: 18.0 per cent, lead; 16.6 per cent, zinc; 6.0 per cent,
iron; 2.5 per cent, manganese; 3.2 per cent, alumina; 2.1 per cent,
lime; 38.5 per cent, insoluble residue; total, 100 per cent. Its
silver content was 19.2 oz. per ton.

The converter charge in this case consisted of 100 parts of
raw slime and 30 parts of gypsum. The converted material
assayed as follows: 16.1 per cent, lead; 14.0 per cent, zinc; 3.6
per cent, sulphur; 5.42 per cent, iron; 2.25 per cent, manganese;
4.10 per cent, alumina; 8.60 per cent, lime; 39.80 per cent, insol-
uble residue; 6.13 per cent, undetermined (oxygen, etc.); total,
100 per cent.; and silver, 17.5 oz. per ton. The increase in weight
of desulphurized ore over that of the raw ore was 11 per cent.
During this run 22 tons of acid were manufactured.

The analysis of the gypsum used in each of the above tests
(at Broken Hill) was as follows: 76.1 per cent. CaSO4, 2H2O;
0.5 per cent. Fe2O3; 4.5 per cent. A12O3; 18.9 per cent, insoluble
residue.

The plant was then put into continuous operation on a mixture
of three parts slime and one of concentrate, desulphurizing down
to 4 per cent. S, and supplying 20 tons of acid per week, and
additions were made to the plant as soon as possible. The acid
made at Broken Hill has been used in connection with the Delprat
process for the concentration of the zinc tailing. At Port Pirie,
works are being erected with capacity for desulphurization of
about 35,000 tons per annum, with an acid output of 10,000 tons.
This acid is to be utilized for the acidulation of phosphate
rock.

The cost of desulphurization of a ton of galena concentrate
by the Carmichael-Bradford process, based on labor at $1.80

LIME-ROASTING OF GALENA

181

182 LEAD SMELTING AND REFINING

per 8 hours, gypsum at $2.40 per 2240 lb., and coal at $8.40 per
2240 lb., is estimated as follows:

0.25 ton of gypsum $0.60

Dehydrating and granulating gypsum 48

Drying mixture of ore and gypsum 12

Converting 24

Spalling sintered material 12

0.01 ton coal 08

Total $1.64

The lime in the sintered product is credited at 12c., making
the net cost $1.52 per ton (2240 lb.) of ore.

The plant required for the Carmichael-Bradford process can
be described with sufficient clearness without drawings, except
the converters. The ore (concentrate, slime, etc.) to be desul-
phurized is delivered at the top of the mill by cars, conveyors,
or other convenient means, and dumped into a bin. Two screw
feeders placed inside the bin supply the mill with ore, uniformly
and as fast as it is required. These feeders deliver the ore into
a chute, which directs it into a vertical dry mixer.

A small bin, on the same level as the ore-bin, receives the crude
gypsum from cars. Thence it is fed automatically to a disin-
tegrator, which pulverizes it finely and delivers it into a storage
bin underneath. This disintegrator revolves at about 1700 r.p.m.
and requires 10 h.p. The body of the machine is cast iron, fitted
with renewable wearing plates (made of hard iron) in the grinding
chamber. The revolving parts consist of a malleable iron disc
in which are fixed steel beaters, faced on the grinding surface
with highly tempered steel. The bin that receives the floured
gypsum contains a screw conveyor similar to those in the ore-
bin, and dumps the material into push conveyors passing into the
dehydrating furnace. They carry the crushed gypsum along at
a speed of about 1 ft. per minute, and allow about 20 ft. to de-
hydrate the gypsum. This speed can, of course, be regulated to
suit requirements.

The dehydrated gypsum runs down a chute into an elevator
boot, and is elevated into a bin which is on the same level as the
ore-bin. This bin also contains a screw conveyor, like that in
the ore-bin. The speed of delivery is regulated to deliver the
right proportion of dehydrated gypsum to the mixer.

LIM&-ROASTING OF GALENA 183

The mixer is of the vertical pattern and receives the sulphide
ore and dehydrated gypsum from the screw feeders. In it are
set two flat revolving cones running at different speeds, thus
ensuring a thorough mixture of the gypsum and ore. The mixed
material drops from the cones upon two baffle plates, and is
wetted just before entering the pug-mill. The pug-mill is a
wrought-iron cylinder of J-in. plate about 2 ft. 6 in. diameter
and 6 or 8 ft. long, and has the mixer fitted to the head. It
contains a 3-ft. wrought-iron spiral with propelling blades, which
forces the plastic mixture through f-in. holes in the cover. The
material comes out in long cylindrical pieces, but is broken up
and formed into marble-shaped pieces on dropping into a revolving
trommel.

The trommel is about 5 ft. long, 2 ft. in diameter at the small
end and about 4 ft. at the large end. It revolves about a wrought-
iron spindle (2J in. diameter) carrying two cast-iron hubs to
which are fitted arms for carrying the conical plate J in. thick.
About 18 in. of the small end of the cone is fitted with wire gauze,
so as to prevent the material as it comes out of the pug-mill from
sticking to it. The trommel is driven by bevel gearing at 20 to
25 r.p.m. The granulated material formed in the trommel is
delivered upon a drying conveyor.

The conveyor consists of hinged wrought-iron plates flanged
at the side to keep the material from running off. It is driven
from the head by gearing, at a speed of 1 ft. per minute, passing
through a furnace 10 ft. long to dry and set the granules of ore
and gypsum. This speed can, of course, be regulated to suit
requirements. The granulated material, after leaving the fur-
nace, is delivered to a single-chain elevator, traveling at a speed
of about 150 ft. per minute. It drops the material into a grass-
hopper conveyor, driven by an eccentric, which distributes the
material over the length of a storage bin. From this bin the
material is directed into the converters by means of the chutes,
which have their bottom ends hinged so as to allow for the raising
of the hood when charging the converters.

The converters are shown in the accompanying engravings,
but they may be of slightly different form from what is shown
therein, i.e., they may be more spherical than conical. They
will have a capacity of about four tons, being 6 ft. in diameter
at the top, 4 ft. in diameter at the false bottom, and about 5 ft.

184

LEAD SMELTING AND REFINING

LIME-ROASTING OF GALENA 185

deep. They are swung on cast-iron trunnions bolted to the
body and turned by means of a hand- wheel and worm (not shown) .
They are carried on strong cast-iron standards fitted with bearings
for trunnions, and all necessary brackets for tilting gear. The
hood has a telescopic funnel which allows it to be raised or lowered,
weights being used to balance it. At the apex of the cone a
damper is provided to regulate the draft. A 4-in. hole in the
pot allows the air from the blast-pipe, 18 in. in diameter, to
enter under the false perforated bottom, the connection between
the two being made by a flexible pipe and coupling. Two Baker
blowers supply the blast for the converters. The material, after
being sintered, is tipped on the floor in front of the converters
and is there broken up to any suitable size, and thence dispatched
to the smelters.

The necessary power for a plant with a capacity of 150 tons
of ore per day will be supplied by a 50-h.p. engine.

THE SAVELSBERG PROCESS

BY WALTER RENTON INGALLS

(December 9, 1905)

There are in use at the present time three processes for the
desulphurization of galena by the new method, which has been
referred to as the " lime-roasting of galena." The Huntington-
Heberlein and the Carmichael-Bradford processes have been pre-
viously described. The third process of this type, which in
some respects is more remarkable than either of the others, is
the invention of Adolf Savelsberg, director of the smeltery at
Ramsbeck, Westphalia, Germany, which is owned by the Akt.
Gesell. f. Bergbau, Blei. u. Zinkhiittenbetrieb zu Stollberg u. in
Westphalen. The process is in use at the Ramsbeck and Stol-
berg lead smelteries of that company. It is described in Ameri-
can patent No. 755,598, issued March 22, 1904 (application filed
Dec. 18, 1903). The process is well outlined in the words of the
inventor in the specification of that patent:

"The desulphurizing of certain ores has been effected by
blowing air through the ore in a chamber, with the object of
doing away with the imperfect and costly process of roasting in
ordinary furnaces; but hitherto it has not been possible satisfac-
torily to desulphurize lead ores in this manner, as, if air be blown
through raw lead ores in accordance with either of the processes
used for treating copper ores, for example, the temperature rises
so rapidly that the unroasted lead ore melts and the air can no
longer act properly upon it, because by reason of this melting
the surface of the ores is considerably decreased, the greater
number of points or extent of surface which the raw ore originally
presented to the action of the oxygen of the air blown through
being lost, and, moreover, the further blowing of air through
the molten mass of ore produces metallic lead and a plumbiferous
slag (in which the lead oxide combines with the gangue) and also
a large amount of light dust, consisting mainly of sublimated
lead sulphide. Huntington and Heberlein have proposed to

186

LIME-ROASTING OF GALENA 187

overcome these objections by adopting a middle course, consisting
in roasting the ores with the addition of limestone for overcoming
the ready fusibility of the ores, and then subjecting them to the
action of the current of air in the chamber; but this process is
not satisfactory, because it still requires the costly previous
operation in a roasting furnace.

"My invention is based on the observation which I have
made that if the lead ores to be desulphurized contain a sufficient
quantity of limestone it is possible, by observing certain precau-
tions, to dispense entirely with the previous roasting in a roasting
furnace, and to desulphurize the ores in one operation by blowing
air through them. I have found that the addition of limestone
renders the roasting of the lead ore unnecessary, because the
limestone produces the following effects:

"The particles of limestone act mechanically by separating
the particles of lead ore from each other in such a way that prem-
ature agglomeration is prevented and the whole mass is loosened
and rendered accessible to air; and, moreover, the limestone
moderates the high reaction temperature resulting from the
burning of the sulphur, so that the liquefaction of the galena,
the sublimation of lead sulphide, and the separation of metallic
lead are avoided. The moderation of the temperature of reaction
is caused by the decomposition of the limestone into caustic
lime and carbon dioxide, whereby a large amount of heat becomes,
latent. Further, the decomposition of the limestone causes chem-
ical reactions, lime being formed, which at the moment of it&
formation is partly converted into sulphate of lime at the expense
of the sulphur contained in the ore, and this sulphate of lime,
when the scorification takes place, is transformed into calcium
silicate by the silicic acid, the sulphuric acid produced thereby
escaping. The limestone also largely contributes to the desul-
phurization of the ore, as it causes the production of sulphuric
acid at the expense of the sulphur of the ore, which sulphuric
acid is a powerful oxidizing agent. If, therefore, a mixture of
raw lead ore and limestone (which mixture must, of course,
contain a sufficient amount of silicic acid for forming silicates)
be introduced into a chamber and a current of air be blown
through the mixture, and at the same time the part of the mixture
which is near the blast inlet be ignited, the combustion of the
sulphur will give rise to very energetic reactions, and sulphurous

188 LEAD SMELTING AND REFINING

acid, sulphuric acid, lead oxide, sulphates and silicates are pro-
duced. The sulphurous acid and the carbon dioxide escape,
while the sulphuric acid and sulphates act in their turn as oxi-
dizing agents on the undecomposed galena. Part of the sulphates
is decomposed by the silicic acid, thereby liberating sulphuric
acid, which, as already stated, acts as an oxidizing agent. The
remaining lead oxide combines finally with the gangue of the
ore and the non- volatile constituents of the flux (the limestone)
to form the required slag. These several reactions commence at
the blast inlet at the bottom of the chamber, and extend grad-
ually toward the upper portion of the charge of ore and limestone.
Liquefaction of the ores does not take place, for although a slag
is formed it is at once solidified by the blowing in of the air, the
passages formed thereby in the hardening slag allowing of the
continued passage therethrough of the air. The final product is
a silicate consisting of lead oxide, lime, silicic acid, and other
constituents of the ore, which now contains but little or no sul-
phur and constitutes a coherent solid mass, which, when broken
into pieces, forms a material suitable to be smelted.

"The quantity of limestone required for the treatment of the
lead ores varies according to the constitution of the ores. It
should, however, amount generally to from 15 to 20 per cent.
As lead ores do not contain the necessary amount of limestone
as a natural constituent, a considerable amount of limestone
must be added to them, and this addition may be made .either
during the dressing of the ores or subsequently.

"For the satisfactory working of the process, the following
precautions are to be observed: In order that the blowing in of
the air may not cause particles of limestone to escape in the
form of dust before the reaction begins, it is necessary to add to
the charge before it is subjected to the action in the chamber a
considerable amount of water — say 5 per cent, or more. This
water prevents the escape of dust, and it also contributes con-
siderably to the formation of sulphuric acid, which, by its oxi-
dizing action, promotes the reaction, and, consequently, also the
desulphurization. It is advisable, in conducting the operation,
not to fill the chamber with the charge at once, but first only
partly to fill it and add to the charge gradually while the chamber
is at work, as by this means the reaction will take place more
smoothly in the mass.

LIME-ROASTING OF GALENA 189

"It is advantageous to proceed as follows: The bottom part
of a chamber of any suitable form is provided with a grate, on
which is laid and ignited a mixture of fuel (coal, coke, or the like)
and pieces of limestone. By mixing the fuel with pieces of lime-
stone the heating power of the fuel is reduced and the grate is
protected, while at the same time premature melting of the
lower part of the charge is prevented; or the grate may be first
covered with a layer of limestone and the fuel be laid thereon ,
and then another layer of limestone be placed on the fuel. On
the material thus placed in the chamber, a uniform charge of
lead ore and limestone — say about 12 in. high — is placed, this
having been moistened as previously explained. Under the in-
fluence of the air-blast and the heat, the reactions hereinbefore
described take place. When the upper surface of the first layer
becomes red-hot, a further charge is laid thereon, and further
charges are gradually introduced as the surface of the preceding
charge becomes red-hot, until the chamber is full. So long as
charges are still introduced a blast of air of but low pressure is
blown through; but when the chamber is filled a larger quantity
of air at a higher pressure is blown through. The scorification
process then takes place, a very powerful desulphurization having
preceded it. During the scorification the desulphurization is
completed.

"When the process is completed, the chamber is tilted and
the desulphurized mass falls out and is broken into small pieces
for smelting."

The drawing on page 190, Fig. 17, shows a side view of the
apparatus used in connection with the process, which will be
readily understood without special description. The dotted lines
show the pot in its emptying position. The series of operations
is clearly illustrated in Figs. 18-20, which are reproduced from
photographs.

This process has now been in practical use at Ramsbeck for
three years, where it is employed for the desulphurization of
galena of high grade in lead, with which are mixed quartzose
silver ore (or sand if no such ore be available), and calcareous and
ferruginous fluxes. A typical charge is 100 parts of lead ore,
10 parts of quartzose silver ore, 10 parts of spathic iron ore, and
19 parts of limestone. A thorough mixture of the components
is essential; afte*- the mixture has been effected, the charge is

190

LEAD SMELTING AND REFINING

thoroughly wetted with about 5 per cent, of water, which is
conceived to play a threefold function in the desulphurizing
operation, namely: (1) preservation of the homogeneity of the
mixture during the blowing; (2) reduction of temperature during
the process; and (3) formation of sulphuric acid in the process,
which promotes the desulphurization of the ore.

The moistened charge is conveyed to the converters, into
which it is fed in thin layers. The converters are hemispherical
cast-iron pots, supported by trunnions on a truck, as shown in

FIG. 17. -Savelsberg Converter.

the accompanying engravings. Except for this method of support,
which renders the pot movable, the arrangement is quite similar
to that which is employed in the Huntington-Heberlein process.
The pots which are now in use at Ramsbeck have capacity for
about 8000 kg. of charge, but it is the intention of the manage-
ment to increase the capacity to 10,000 or 12,000 kg. Previously,
pots of only 5000 kg. capacity were employed. Such a pot
weighed 1300 kg., exclusive of the truck. The air-blast was
about 7 cu. m. (247.2 cu. ft.) per min., beginning at a pressure

Fig. 20.— Converter in Position for Blowing.

LIME-ROASTING OF GALENA 191

of 10 to 20 cm. of water (2f to 4^ oz.) and rising to 50 to 60 cm.
(11| to 13^ oz.) when the pot was completely filled with charge.
The desulphurization of a charge is completed in 18 hours. A
pot is attended by one man per shift of 12 hours; this is only the
attention of the pot proper, the labor of conveying material to it
and breaking up the desulphurized product being extra. One
man per shift should be able to attend to two pots, which is the
practice in the Huntington-Heberlein plants.

When the operation in the pot is completed, the latter is turned
on its trunnions, until the charge slides out by gravity, which it
does as a solid cake. This is caused to fall upon a vertical bar,
which breaks it into large pieces. By wedging and sledging these
are reduced to lumps of suitable size for the blast furnace. When
the operation has been properly conducted the charge is reduced
to about 2 or 3 per cent, sulphur. It is expected that the use
of larger converters will show even more favorable results in this
particular.

As in the Huntington-Heberlein and Carmichael-Bradford
processes, one of the greatest advantages of the Savelsberg
process is the ability to effect a technically high degree of desul-
phurization with only a slight loss of lead and silver, which is of
course due to the perfect control of the temperature in the process.
The precise loss of lead has not yet been determined, but in the
desulphurization of galena containing 60 to 78 per cent, lead,
the loss of lead is probably not more than 1 per cent. There
appears to be no loss of silver.

The process is applicable to a wide variety of lead-sulphide
ores. The ore treated at Ramsbeck contains 60 to 78 per cent,
lead and about 15 per cent, of sulphur, but ore from Broken Hill,
New South Wales, containing 10 per cent, of zinc has also been
treated. A zinc content up to 7 or 8 per cent, in the ore is no
drawback, but ores carrying a higher percentage of zinc require
a larger addition of silica and about 5 per cent, of iron ore in
order to increase the fusibility of the charge. The charge ordi-
narily treated at Ramsbeck is made to contain about 11 per cent,
of silica. The presence of pyrites in the ore is favorable to the
desulphurization. Dolomite plays the same part in the process
that limestone does, but is of course less desirable, in view of the
subsequent smelting in the blast furnace. The ore is best crushed
to about 3 mm. size, but good results have been obtained with

192 LEAD SMELTING AND REFINING

ore coarser in size than that. However, the proper size is some-
what dependent upon the character of the ore. The blast pressure
required in the converter is also, of course, somewhat dependent
upon the porosity of the charge. Fine slimes are worked up by
mixture with coarser ore.

In making up the charge, the proportion of limestone is not
varied much, but the proportions of silica and iron must be
carefully modified to suit the ore. Certain kinds of ore have a
tendency to remain pulverulent, or to retain balls of unsintered,
powdered material. In such cases it is necessary to provide more
fusible material in the charge, which is done by varying the
proportions of silica and iron. The charge must, moreover, be
prepared in such a manner that overheating, and consequently
the troublesome fusion of raw galena, will be avoided.

The essential difference between the Huntington-Heberlein
and Savelsberg processes is the use in the former of a partially
desulphurized ore, containing lime and sulphate of lime; and the
use in the latter of raw ore and carbonate of lime. It is claimed
that the latter, which loses its carbon dioxide in the converter,
necessarily plays a different chemical part from that of quicklime
or gypsum. Irrespective of the reactions, however, the Savelsberg
process has the great economic advantage of dispensing with the
preliminary roasting of the Huntington-Heberlein process, where-
fore it is cheaper both in first cost of plant and in operation.

THE LIME-ROASTING OF GALENA1

BY WALTER RENTON INGALLS

During the last two years, and especially during the last six
months, a number of important articles upon the new methods
for the desulphurization of galena have been published in the
technical periodicals, particularly in the Engineering and Mining
Journal and in Metallurgie. I proposed for these methods the
type-name of " lime-roasting of galena," as a convenient metal-
lurgical classification,2 and this term has found some acceptance.
The articles referred to have shown the great practical importance
of these new processes, and the general recognition of their
metallurgical and commercial value, which has already been
accorded to them. It is my present purpose to review broadly
the changes developed by them in the metallurgy of lead, in
which connection it is necessary to refer briefly to the previous
state of the art.

The elimination of the sulphur content of galena has been
always the most troublesome part of the smelting process, being
both costly in the operation and wasteful of silver and lead.
Previous to the introduction of the Huntington-Heberlein process
at Pertusola, Italy, it was effected by a variety of methods. In
the treatment of non-argentiferous galena concentrate, the smelt-
ing was done by the roast-reduction method (roasting in rever-
beratory furnace and smelting in blast furnace) ; the roast-reaction
method, applied in reverberatory furnaces; and the roast-reaction
method, applied in Scotch hearths.3 Precipitation smelting,
simple, had practically gone out of use, although its reactions
enter into the modern blast-furnace practice, as do also those of
the roast-reaction method.

1 A paper presented before the American Institute of Mining Engineers,
July, 1906.

2 Engineering and Mining Journal, Sept. 2, 1905.

3 This term is inexact, because the hearths employed in the United
States are not strictly "Scotch hearths," but they are commonly known as
such, wherefore my use of the term.

193

194 LEAD SMELTING AND REFINING

In the treatment of argentiferous lead ores, a combination of
the roast-reduction, roast-reaction and precipitation methods
had been developed. Ores low in lead were still roasted, chiefly
in hand-worked reverberatories (the mechanical furnaces not
having proved well adapted to lead-bearing ores), while the high
loss of lead and silver in sinter- or slag-roasting of rich galenas
had caused those processes to be abandoned, and such ores were
charged raw into the blast furnace, the part of their sulphur
which escaped oxidation therein reappearing in the form of
matte. In the roast-reduction smelting of galena alone, how-
ever, there was no way of avoiding the roasting of the whole, or
at least a very large percentage of the ore, and in this roasting
the ore had necessarily to be slagged or sintered in order to elim-
inate the sulphur to a satisfactory extent. This is exemplified
in the treatment of the galena concentrate of southeastern Mis-
souri at the present time.

Until the two new Scotch-hearth plants at Alton and Collins-
ville, 111., were put in operation, the three processes of smelting
the southeastern Missouri galena were about on an equal footing.
Their results per ton of ore containing 65 per cent, lead were
approximately as follows l:

METHOD

COST

EXTRAC-
TION

Reverberatory

$6.50-7.00

90-92%

Scotch hearth . .

5 75-6 50

87-88%

Roast-reduction ...

600-700

90-92%

The new works employ the Scotch-hearth process, with bag-
houses for the recovery of the fume, which previously was the
weak point of this method of smelting.2 This improvement led
to a large increase in the recovery of lead, so that the entire
extraction is now approximately 98 per cent, of the content of
the ore, while on the other hand the cost of smelting per ton of

1 Percentages of lead in Missouri practice are based on the wet assay;
among the silver-lead smelters of the West the fire assay is still generally
employed.

2 This improvement did not originate at either Alton or Collinsville. It
had previously been hi use at the works of the Missouri Smelting Company
at Cheltenham, St. Louis, but the idea originated from the practice of the
Picher Lead Company, of Joplin, Mo.

LIME-ROASTING OF GALENA 195

ore has been reduced through the increased size of these plants
and the introduction of improved means for handling ore and
material. The practice of these works represents the highest
efficiency yet obtained in this country in the smelting of high-
grade galena concentrate, and probably it cannot be equaled
•even by the Huntington-Heberlein and similar processes. The
Scotch-hearth and bag-house process is therefore the one of the
older methods of smelting which will survive.

In the other methods of smelting, a large proportion of the
cost is involved in the roasting of the ore, which amounts in
hand- worked reverberatory furnaces to $2 to $2.50 per ton.
Also, the larger proportion of the loss of metal is suffered in the
roasting of the ore, this loss amounting to from 6 to 8 per cent,
of the metal content of such ore as is roasted. The loss of lead
in the combined process of treatment depends upon the details
of the process. The chief advantage of lime-roasting in the treat-
ment of this class of ore is in the higher extraction of metal which
it affords. This should rise to 98 per cent. That figure has
been, indeed, surpassed in operations on a large scale, extending
over a considerable period.

In the treatment of the argentiferous ores of the West different
conditions enter into the consideration. In the working of those
ores, the present practice is to roast only those which are low in
lead, and charge raw into the blast furnace the rich galenas.
The cost of roasting is about $2 to $2.50 per ton; the cost of
smelting is about $2.50 per ton. On the average about 0.4 ton
of ore has to be roasted for every ton that is smelted. The cost
of roasting and smelting is therefore about $3.50 per ton. In
good practice the recovery of silver is about 98 per cent, and of
lead about 95 per cent., reckoned on basis of fire assays.

In treatment of these ores, the lime-roasting process offers
several advantages. It may be performed at less than the cost
of ordinary roasting.1 The loss of silver and lead during the
roasting is reduced to insignificant proportion. The sulphide
fines which must be charged raw into the blast furnace are elimi-
nated, inasmuch as they can be efficiently desulphurized in the
lime-roasting pots without significant loss; all the ore to be
smelted in the blast furnace can be, therefore, delivered to it in
lump form, whereby the speed of the blast furnace is increased
1 This refers especially to the Savelsberg process.

196 LEAD SMELTING AND REFINING

and the wind pressure required is decreased. Finally, the per-
centage of sulphur in the charge is reduced, producing a lower
matte-fall, or no matte-fall whatever, with consequent saving
in expense of retreatment. In the case of a new plant, the first
cost of construction and the ground-space occupied are materially
reduced. Before discussing more fully the extent and nature of
these savings, it is advisable to point out the differences among
the three processes of lime-roasting that have already come into
practical use.

In the Huntington-Heberlein process, the ore is mixed with
suitable proportions of limestone and silica (or quart zose ore) and
is then partially roasted, say to reduction of the sulphur to one
half. The roasting is done at a comparatively low temperature,
and the loss of metals is consequently small. The roasted ore is
dampened and allowed to cool. It is then charged into a hemi-
spherical cast-iron pot, with a movable hood which covers the
top and conveys off the gases. There is a perforated grate in
the bottom of the pot, on which the ore rests, and air is introduced
through a pipe entering the bottom of the pot, under the grate.
A small quantity of red-hot calcines from the roasting furnaces
is thrown on the grate to start the reaction; a layer of cold,
semi-roasted ore is put upon it, the air blast is turned on and
reaction begins, which manifests itself by the copious evolution
of sulphur fumes. These consist chiefly of sulphur dioxide, but
they contain more or less trioxide, which is evident from the
solution of copperas that trickles from the hoods and iron smoke-
pipes, wherein the moisture condenses. As the reaction pro-
gresses, and the heat creeps up, more ore is introduced, layer by
layer, until the pot is full. Care is taken by the operator to
compel the air to pass evenly and gently through the charge,
wherefore he is watchful to close blow-holes which develop in it.
At the end of the operation, which may last from four to eighteen
hours, the ore becomes red-hot at the top. The hood is then
pushed up, and the pot is turned on its trunnions, by means of
a hand-operated wheel and worm-gear, until the charge slides out,
which it does as a solid, semi-fused cake. The pot is then turned
back into position. Its design is such that the air-pipe makes
automatic connection, a flanged pipe cast with the pot settling
upon a similiarly flanged pipe communicating with the main, a
suitable gasket serving to make a tight joint. The pots are set

LIME-ROASTING OF GALENA 197

at an elevation of about 12 ft. above the ground, so that when
the charge slides out the drop will break it up to some extent,
and it is moreover caused to fall on a wedge, or similar contriv-
ance, to assist the breakage. After cooling it is further broken
up to furnace size by wedging and sledging; the lumps are forked
out, and the fines screened and returned to a subsequent charge
for completion of their desulphurization.

The Savelsberg process differs from the Huntington-Heberlein
in respect to the preliminary roasting, which in the Savelsberg
process is omitted, the raw ore, mixed with limestone and silica,
being charged directly into the converter. The Savelsberg con-
verter is supported on a truck, instead of being fixed in position,
but otherwise its design and management are quite similar to
those of the Huntington-Heberlein converter. In neither case
are there any patents on the converters. The patents are on
the processes. In view of the litigation that has already been
commenced between their respective owners, it is interesting to
examine the claims.

The Huntington-Heberlein patent (U. S. 600,347, issued
March 8, 1898, applied for Dec. 9, 1896) has the following claims:

1. The herein-described method of oxidizing sulphide ores of
lead preparatory to reduction to metal, which consists in mixing
with the ore to be treated an oxide of an alkaline-earth metal,
such as calcium oxide, subjecting the mixture to heat in the
presence of air, then reducing the temperature and finally passing
air through the mass to complete the oxidation of the lead, sub-
stantially as and for the purpose set forth.

2. The herein-described method of oxidizing sulphide ores of
lead preparatory to reduction to metal, which consists in mixing
calcium oxide or other oxide of an alkaline-earth metal with the
ore to be treated, subjecting the mixture in the presence of air
to a bright-red heat (about 700 deg. C.), then cooling down the
mixture to a dull-red heat (about 500 deg. C.), and finally forcing
air through the mass until the lead ore, reduced to an oxide, fuses,
substantially as set forth.

3. The herein-described method of oxidizing lead sulphide in
the preparation of the same for reduction to metal, which consists
in subjecting the sulphide to a high temperature in the presence of
an oxide of an alkaline-earth metal, such as calcium oxide, and oxy-
gen, and then lowering the temperature substantially as set forth.

198 LEAD SMELTING AND REFINING

Adolf Savelsberg, in U. S. patent 755,598 (issued March 22,
1904, applied for Dec. 18, 1903) claims:

1. The herein-described process of desulphurizing lead ores,
which consists in mixing raw ore with limestone and then sub-
jecting the mixture to the simultaneous application of heat and
a current of air in sufficient proportions to substantially complete
the desulphurization in one operation, substantially as described.

2. The herein-described process of desulphurizing lead ores,
which process consists in first mixing the ores with limestone,
then moistening the mixture, then filling it without previous
roasting into a chamber, then heating it and treating it by a current
of air, as and for the purpose described.

3. The herein-described process of desulphurizing lead ores,
which consists in mixing raw ores with limestone, then filling the
mixture into a chamber, then subjecting the mixture to the
simultaneous application of heat and a current of air in sufficient
proportions to substantially complete the desulphurization in one
operation, the mixture being introduced into the chamber in
partial charges introduced successively at intervals during the
process, substantially as described.

4. The herein-described process of desulphurizing lead ores,
then moistening the mixture, then filling it without previous
roasting into a chamber, then heating it and treating it by a
current of air, the mixture being introduced into the chamber
in partial charges introduced successively at intervals during the
process, as and for the purpose described.

5. The herein-described process of desulphurizing lead ores,
which process consists in first mixing the ores with sufficient
limestone to keep the temperature of the mixture below the
melting-point of the ore, then filling the mixture into a chamber,
then heating said mixture and treating it with a current of air,
as and for the purpose described.

6. The herein-described process of desulphurizing lead ores,
which process consists in first mixing the ores with sufficient
limestone to mechanically separate the particles of galena suffi-
ciently to prevent fusion, and to keep the temperature below the
melting-point of the ore by the liberation of carbon dioxide, then
filling the mixture into a chamber, then heating said mixture
and treating it with a current of air, as and for the purpose de-
scribed.

LIME-ROASTING OF GALENA 199

The Carmichael-Bradford process differs from the Savelsberg
by the treatment of the raw ore mixed with gypsum instead of
limestone, and differs from the Huntington-Heberlein both in
respect to the use of gypsum and the omission of the preliminary
roasting. The Carmichael-Bradford process has not been threat-
ened with litigation, so far as I am aware. The claims of its
original patent read as follows 1 :

1. The process of treating mixed sulphide ores, which consists
in mixing with said ores a sulphur compound of a metal of the
alkaline earths, starting the reaction by heating the same, thereby
oxidizing the sulphide and reducing the sulphur compound of
the alkali metal, passing a current of air to oxidize the reduced
sulphide compound of the metal of the alkalies preparatory to
acting upon a new charge of sulphide ores, substantially as and
for the purpose set forth.

2. The process of treating mixed sulphide ores, which consists
in mixing calcium sulphate with said ores, starting the reaction
by means of heat, thereby oxidizing the sulphide ores, liberating
sulphurous-acid gas and converting the calcium sulphate into
calcium sulphide and oxidizing the calcium sulphide to sulphate
preparatory to treating a fresh charge of sulphide ores, substan-
tially as and for the purpose set forth.

The process described by W. S. Bayston, of Melbourne (Aus-
tralian patent No. 2862) , appears to be identical with that of
Savelsberg.

Irrespective of the validity of the Savelsberg and Carmichael-
Bradford patents, and without attempting to minimize the
ingenuity of their inventors and the importance of their discov-
eries, it must be conceded that the merit for the invention and
introduction of lime-roasting of galena belongs to Thomas Hunt-
ington and Ferdinand Heberlein. The former is an American,
and this is the only claim that the United States can make to a
share in this great improvement in the metallurgy of lead. It is
to be regretted, moreover, that of all the important lead-smelting
countries in the world, America has been the most backward in
adopting it.

The details of the three processes and the general results
accomplished by them have been rather fully described in a
series of articles recently published in the Engineering and Mining
* A. D. Carmichael, U. S. patent No. 705,904, July 29, 1902.

200 LEAD SMELTING AND REFINING

Journal. There has been, however, comparatively little discus-
sion as to costs; and unfortunately the data available for analysis
are extremely scanty, due to the secrecy with which the Hunt-
ington-Heberlein process, the most extensively exploited of the
three, has been veiled. Nevertheless, I may attempt an approx-
imate estimation of the various details, taking the Huntington-
Heberlein process as the basis.

The ore, limestone and silica are crushed to pass a four-mesh
screen. This is about the size to which it would be necessary to
crush as preliminary to roasting in the ordinary way, wherefore
the only difference in cost is the charge for crushing the limestone
and silica, which in the aggregate may amount to one-sixth of
the weight of the raw sulphide and may consequently add 2 to
2.5c. to the cost of treating a ton of ore. The mixing of ore and
fluxes may be costly or cheap, according to the way of doing it.
If done in a rational way it ought not to cost more than lOc. per
ton of ore, and may come to less. The delivery of the ore from
the mixing-house to the roasting furnaces ought to be done
entirely by mechanical means, at insignificant cost.

The Heberlein roasting furnace, which is used in connection
with the H.-H. process, is simply an improvement on the old
Brunton calciner — a circular furnace, with revolving hearth.
The construction of this furnace, according to American designs,
is excellent. The hearth is 26 ft. in diameter; it is revolved at
slow speed and requires about 1.5 h.p. A flange at the periphery
of the hearth dips into sand in an annular trough, thus shutting
off air from the combustion chamber, except through the ports
designed for its admittance. The mechanical construction of
the furnace is workmanlike, and the mechanism under the hearth
is easy of access and comfortably attended to.

A 26-ft. furnace roasts about 80,000 Ib. of charge per 24 hours.
In dealing with an ore containing 20 to 22 per cent, of sulphur,
the latter is reduced to about 10 to 11 per cent., the consumption
of coal being about 22.5 per cent, of the weight of the charge.
The hearth efficiency is about 150 Ib. per sq. ft., which in com-
parison with ordinary roasting is high. The coal consumption,
however, is not correspondingly low. Two furnaces can be man-
aged by one man per 8-hour shift. On the basis of 80 tons of
charge ore per 24 hours, the cost of roasting should be approxi-
mately as follows:

LIME-ROASTING OF GALENA 201

Labor— 3 men at $2.50 $ 7.50

Coal — 18 tons at $2 36.00

Power 3.35

Repairs 3.35

Total $50.20 = 63c. per ton.

In the above estimate repairs have been reckoned at the
same figure as is experienced with Bruckner cylinders, and the
cost of power has been allowed for with fair liberality. The
estimated cost of 63c. per ton is comparable with the $1.10 to
$1.45 per ton, which is the result of roasting in Bruckner cylinders
in Colorado, reducing the ore to 4.5-6 per cent, sulphur.

The Heberlein furnace is built up to considerable elevation
above the ground level, externally somewhat resembling the
Pearce turret furnace. This serves two purposes: (1) it affords
ample room under the hearth for attention to the driving mecha-
nism; and (2) it enables the ore to be discharged by gravity into
suitable hoppers, without the construction of subterranean gang-
ways. The ore discharges continuously from the furnace, at
dull-red heat, into a brick bin, wherein it is cooled by a water-
spray. Periodically a little ore is diverted into a side bin, in
which it is kept hot for starting a subsequent charge in the con-
verter.

The cooled ore is conveyed from the receiving bins at the
roasting furnaces to hopper-bins above the converters. If the
tramming be done by hand the cost, with labor at 25c. per hour,
may be approximately 12.5c. per ton of ore, but this . should be
capable of considerable reduction by mechanical conveyance.

The converters are hemispherical pots of cast iron, 9 ft. in
diameter at the top, and about 4 ft. in depth. They are provided
with a circular, cast-iron grate, which is f in. thick and 6 ft. in
diameter and is set and secured horizontally in the pot. This
grate is perforated with holes } in. in diameter, 2 in. apart, center
to center, and is similar to the Wetherill grate employed in zinc
oxide manufacture. The pot itself is about 2J in. thick at the
bottom, thinning to about 1J in. at the rim. It is supported on
trunnions and is geared for convenient turning by hand. The
blast pipe which enters the pot at the bottom is 6 in. in diameter.

Two roasting furnaces and six converters are rated nominally
as a 90-ton plant. This rating is, however, considerably in excess
of the actual capacity, at least on certain ores. The time required

202 LEAD SMELTING AND REFINING

for desulphurization in the converter apparently depends a good
deal upon the character of the ore. The six converters may be
arranged in a single row, or in two rows of three in each. They
are set so that the rim of the pot, when upright, is about 12 ft.
above the ground level. A platform gives access to the pots.
One man per shift can attend to two pots. His work consists in
charging them, which is done by gravity, spreading out the
charge evenly in the pot, closing any blow-holes which may
develop, and at the end of the operation raising the hood (which
covers the pot during the operation) and dumping the pot.
The work is easy. The conditions under which it is done are
comfortable, both as to temperature and atmosphere. Reports
have shown a great reduction in liability to lead-poisoning in
the works where the H.-H. process has been introduced.

A new charge is started by kindling a small wood or coal fire
on the grate, then throwing in a few shovelfuls of hot calcines,
and finally dropping in the regular charge of damp ore (plus the
fluxes previously referred to). The charge is introduced in stages,
successive layers being dropped in and spread out as the heat
rises. At the beginning the blast is very low — about 2 oz. It
is increased as the hight of the ore in the pot rises, finally attain-
ing about 16 oz. The operation goes on quietly, the smoke
rising from the surface evenly and gently, precisely as in a well-
running blast furnace. While the charge is still black on top,
the hand can be held with perfect comfort, inside of the hood,
immediately over the ore. This explains, of course, why the
volatilization of silver and lead is insignificant. There is, more-
over, little or no loss of ore as dust, because the ore is introduced
damp, and the passage of the air through it is at low velocity.
In the interior of the charge, however, there is high temperature
(evidently much higher than has been stated in some descrip-
tions), as will be shown further on. The conditions in this respect
appear to be analogous to those of the blast furnace, which,
though smelting at a temperature of say 1200 deg. C. at the
level of the tuyeres, suffers only a slight loss of silver and lead
by volatilization.

At the end of the operation in the H.-H. pot, the charge is
dull red at the top, with blow-holes, around which the ore is
bright red. Imperfectly worked charges show masses of well-
fused ore surrounded by masses of only partially altered ore,

LIME-ROASTING OF GALENA 203

a condition which may be ascribed to the irregular penetration
of air through the charge, affording good evidence of the impor-
tant part which air plays in the process. A properly worked
charge is tipped out of the pot as a solid cake, which in falling to
the ground breaks into a few large pieces. As they break, it
appears that the interior of the charge is bright red all through,
and there is a little molten slag which runs out of cavities, pre-
sumably spots where the chemical action has been most intense.
When cold, the thoroughly desulphurized material has the ap-
pearance of slag-roasted galena. Prills of metallic lead are visible
in it, indicating reaction between lead sulphide and lead sulphate.

The columns of the structure supporting the pots should be
of steel, since fragments of the red-hot ore dumped on the ground
are likely to fall against them. To hasten the cooling of the
ore, water is sometimes played on it from a hose. This is bad,
since some is likely to splash into the still inverted pot, leading
to cracks. The cracked pots at certain works appear to be due
chiefly to this cause, in the absence of which the pots ought to
last a long time, inasmuch as the conditions to which they are
subjected during the blowing process are not at all severe. When
the ore is sufficiently cold it is further broken up, first by driving
in wedges, and finally by sledging down to pieces of orange size,
or what is suitable for the blast furnace. These are forked out,
leaving the fine ore, which comes largely from the top of the
charge and is therefore only partially desulphurized. The fines
are, therefore, re-treated with a subsequent charge. The quantity
is not excessive; it may amount to 7 or 8 per cent, of the charge.

The breaking up of the desulphurized ore is one of the prob-
lems of the process, the necessity being the reduction of several
large pieces of fused, or semi-fused, material weighing two or
three tons each. When done by hand only, as is usually (per-
haps always) the practice, the operation is rather expensive.
It would appear, however, to be not a difficult matter to devise
some mechanical aids for this process — perhaps to make it
entirely mechanical. When done by hand, a 6-pot plant re-
quires 6 men per shift sledging and forking. With 8-hour
shifts, this is 18 men for the breaking of about 60 tons of material,
which is about 3J tons per man per 8 hours. With labor at
25c. per hour, the cost of breaking the fused material comes to
60c. per ton. It may be remarked, for comparison, that in

204 LEAD SMELTING AND REFINING

breaking ore as it ordinarily comes, coarse and fine together, a
good workman would normally be expected to break 5 to 5.5
tons in a shift of 8 hours.

The ordinary charge for the standard converter is about
8 tons (16,000 Ib.) of an ore weighing 166 Ib. per cu. ft. With
a heavier ore, like a high-grade galena, the charge would weigh
proportionately more. The time of working off a charge is
decidedly variable. Accounts of the operation of the process in
Australia tell of charge- workings in 3 to 5 hours, but this
does not correspond with the results reported elsewhere, which
specify times of 12 to 18 hours. Assuming an average of 16
hours, which was the record of one plant, six converters would
have capacity for about 72 tons of charge per 24 hours, or about
58 tons of ore, the ratio of ore to flux being 4: 1. The loss in
weight of the charge corresponds substantially to the replacement
of sulphur by oxygen, and the expulsion of carbon dioxide. The
finished charge contains on the average from 3 to 5 per cent,
sulphur. This is about the same as the result achieved in good
practice in roasting lead-bearing ores in hand-worked reverbera-
tory furnaces, but curiously the H.-H. product, in some cases at
least, does not yield any matte, to speak of, in the blast furnace;
the product delivered to the latter being evidently in such condi-
tion that the remaining sulphur is almost completely burned off
in the blast furnace. This is an important saving effected by
the process. In calculating the value of an ore, sulphur is
commonly debited at the rate of 25c. per unit, which represents
approximately the cost of handling and reworking the matte
resulting from it. The practically complete elimination of matte-
fall rendered possible by the H.-H. process may not be, however,
an unmixed blessing. There may be, for example, a small for-
mation of lead sulphide which causes trouble in the crucible and
lead-well, and results in furnace difficulties and the presentation
of a vexatious between-product.

It may now be attempted to summarize the cost of the con-
verting process. Assuming the case of an ore assaying lead, 50 per
cent.; iron, 15; sulphur, 22; silica, 8, and alumina, etc., 5, let it be
supposed that it is to be fluxed with pure limestone and pure
quartz, with the aim to make a slag containing silica, 30; ferrous
oxide, 40; and lime, 20 per cent. A ton of ore will make, in
round numbers, 1000 Ib. of slag, and will require 344 Ib. of lime-

LIME-ROASTING OF GALENA

205

stone and 130 Ib. of quartz, or we may say roughly one ton of
flux must be added to four tons of ore, wherefore the ore will
constitute 80 per cent, of the charge. In reducing the charge
to 3 per cent, sulphur it will lose ultimately through expulsion
of sulphur and carbon dioxide (of the limestone) about 20 per cent,
in weight, wherefore the quantity of material to be smelted in
the blast furnace will be practically equivalent to the raw sulphide
ore in the charge for the roasting furnaces; but in the roasting
furnace the charge is likely to gain weight, because of the forma-
tion of sulphates. Taking the charge, which I have assumed
above, and reckoning that as it comes from the roasting furnace
it will contain 10 per cent, sulphur, all in the form of sulphate,
either of lead or of lime, and that the iron be entirely converted
to ferric oxide, in spite of the expulsion of the carbon dioxide of
the limestone and the combustion of a portion of the sulphur of
the ore as sulphur dioxide, the charge will gain in weight in the
ratio of 1 : 1.19. This, however, is too high, inasmuch as a portion
of the sulphur will remain as sulphide while a portion of the iron
may be as ferrous oxide. The actual gain in weight will conse-
quently be probably not more than one-tenth. The following
theoretical calculation will illustrate the changes:

RAW CHARGE

SEMI-ROASTED CHARGE

FINISHED CHARGE

Ore
Flux

1000 Ib. Pb

Ore
Flux

1154 Ib. PbO
428 Ib. Fe2O3
160 Ib. SiO2
100 Ib. Al2O3,etc.
300 Ib S

Ore
Flux

11541b.PbO..
4281b.Fe2O3(?)
160 Ib. SiO2 ....
100 Ib. Al2O3,etc.
68 Ib S

300 Ib. Fe

160 Ib. SiO2
100 Ib. Al2O3,etc.
440 Ib S

130 Ib. SiO2...
344 Ib. CaCO3....

130 Ib SiO2

1301b SiO2

193 Ib CaO

193 Ib.CaO

450 Ib.O

2474 Ib.

2915 Ib.
10% S.

2233 Ib.

3%S.

Ratios:

2474: 2915 ::!:!. 18.
2915: 2233 ::l:0.76f.
2474: 2233:: 1:0.90.

It may be assumed that for every ton of charge (containing
about 80 per cent, of ore) there will be 1.1 ton of material to go
to the converter, and that the product of the latter will be 0.&
of the weight of the original charge of raw material.

206 LEAD SMELTING AND REFINING

Each converter requires 400 cu. ft. of air per minute. The
blast pressure is variable, as different pots are always at different
stages of the process, but assuming the maximum of 16 oz. pres-
sure, with a blast main of sufficient diameter (at least 15 in.)
and the blower reasonably near the battery of pots, the total
requirement is 21 h.p. The cost of converting will be approxi-
mately as follows:

Labor, 3 foremen at $3.20 $ 9.60

" 9 men at $2.50 22.50

Power, 21 h.p. at 30c 6.30

Supplies, repairs and renewals 5.00

Total $43.40= 60c. per ton of charge.

The cost of converting is, of course, reduced directly as the
time is reduced. The above estimate is based on unfavorable
conditions as to time required for working a charge.

The total cost of treatment, from the initial stage to the
delivery of the desulphurized ore to the blast furnaces, will be,
per 2000 Ib. of charge, approximately as follows:

Crushing 1.0 ton at lOc $0.10

Mixing 1.0 ton at lOc 10

Roasting 1.0 ton at 63c 63

Delivering 1.1 ton to converters at 12c 13

Converting 1.1 ton at 60c 66

Breaking 0.9 ton at 60c 54

Total $2.16

The cost per ton of ore will be 2.16 -^ 0.80 = $2.70. Making
allowance for the crushing of the ore, which is not ordinarily
included in the cost of roasting, and possibly some overestimates,
it appears that the cost of desulphurization by this method,
under the conditions assumed in this paper, is rather higher than
in good practice with ordinary hand-worked furnaces, but it is
evident that the cost can be reduced to approximately the same
figure by introduction of improvements, as for example in break-
ing the desulphurized ore, and by shortening the time of con-
verting, which is possible in the case of favorable ores. The chief
advantage must be, however, in the further stage of the smelting.
As to this, there is the evidence that the Broken Hill Proprietary
Company was able to smelt the same quantity of ore in seven

LIME-ROASTING OF GALENA 207

furnaces, after the introduction of the Huntington-Heberlein
process, that formerly required thirteen. A similar experience
is reported at Friedrichshutte, Silesia.

This increase in the capacity of the blast furnace is due to
three things: (1) In delivering to the furnace a charge containing
a reduced percentage of fine ore, the speed of the furnace is
increased, i.e., more tons of ore can be smelted per square foot
of hearth area. (2) There is less roasted matte to go into the
charge. (3) Under some conditions the percentage of lead in
the charge can be increased, reducing the quantity of gangue
that must be fluxed.

It is difficult to generalize the economy that is effected in
the blast-furnace process, since this must necessarily vary within
wide limits because of the difference in conditions. An increase
of 60 to 100 per cent, in blast-furnace capacity does not imply a
corresponding reduction in the cost of smelting. The fuel con-
sumption per ton of ore remains the same. There is a saving in
the power requirements, because the smelting can be done with
a lower blast pressure; also, a saving in the cost of reworking
matte. There will, moreover, be a saving in other labor, in so
far as portions thereof are not already performed at the minimum
cost per ton. The net result under American conditions of
silver-lead smelting can only be determined closely by extensive
operations. That there will be an important saving, however,
there is no doubt.

The cost of smelting a ton of charge at Denver and Pueblo,
exclusive of roasting and general expense, is about $2.50, of
which about $0.84 is for coke and $1.66 for labor, power and
supplies. General expense amounts to about $0.16 additional.
If it should prove possible to smelt in a given plant 50 per cent,
more ore than at present without increase in the total expense,
except for coke, the saving per ton of charge would be 70c. That
is not to be expected, but the half of it would be a satisfactory
improvement. With respect to sulphur in the charge, the cost
is commonly reckoned at 25c. per unit. As compared with a
charge containing 2 per cent, of sulphur there would be a saving
rising toward 50c. per ton as the maximum. It is reasonable to
reckon, therefore, a possible saving of 75c. per ton of charge in
silver-lead smelting, no saving in the cost of roasting, and an
increase of about 3 per cent, in the extraction of lead, and per-

208 LEAD SMELTING AND REFINING

haps 1 per cent, in the extraction of silver, as the net results of
the application of the Huntington-Heberlein process in American
silver-lead smelting.

On a charge averaging 12 per cent, lead and 33 oz. silver per
ton, an increase of 3 per cent, in the extraction of lead and
1 per cent, in the extraction of silver would correspond to 25c.
and 35c. respectively, reckoning lead at 3.5c. per lb., and silver
at 60c. per oz. In this, however, it is assumed that all lead-
bearing ores will be desulphurized by this process, which prac-
tically will hardly be the case. A good deal of pyrites, containing
only a little lead, will doubtless continue to be roasted in Bruckner
cylinders, and other mechanical furnaces, which are better
adapted to the purpose than are the lime-roasting pots. More-
over, a certain proportion of high-grade lead ore, which is now
smelted raw, will be desulphurized outside of the furnace, at
additional expense. It is comparatively simple to estimate the
probable benefit of the Huntington-Heberlein process in the case
of smelting works which treat principally a single class of ore,
but in such works as those in Colorado and Utah, which treat a
wide variety of ores, we must anticipate a combination process,
and await results of experience to determine just how it will
work out. It should be remarked, moreover, that my estimates
do not take into account the royalty on the process, which is an
actual debit, whether it be paid on a tonnage basis or be com-
puted in the form of a lump sum for the license to its use.

However, in view of the immense tonnage of ore smelted
annually for the extraction of silver and lead, it is evident that
the invention of lime-roasting by Huntington and Heberlein was
an improvement of the first order in the metallurgy of lead.

In the case of non-argentiferous galena, containing 65 per
cent, of lead (as in southeastern Missouri), comparison may be
made with the slag-roasting and blast-furnace smelting of the
ore. Here, no saving in cost of roasting may be reckoned and no
gain in the speed of the blast furnaces is to be anticipated. The
only savings will be in the increase in the extraction of lead
from 92 to 98 per cent., and the elimination of matte-roasting,
which latter may be reckoned as amounting to 50c. per ton of
ore. The extent of the advantage over the older method is so
clearly apparent that it need not be computed any further. In
comparison with the Scotch-hearth bag-house method of smelting,

LIME-ROASTING OF GALENA 209

however, the advantage, if any, is not so certain. That method
already saves 98 per cent, of the lead, and on the whole is prob-
ably as cheap in operation as the Huntington-Heberlein could be
under the same conditions. The Huntington-Heberlein method
has replaced the old roast-reaction method at Tarnowitz, Silesia,
but the American Scotch-hearth method as practised near St.
Louis is likely to survive.

A more serious competitor will be, however, the Savelsberg
process, which appears to do all that the Huntington-Heberlein
process does, without the preliminary roasting. Indeed, if the
latter be omitted (together with its estimated expense of 63c.
per ton of charge, or 79c. per ton of ore), all that has been said
in this paper as to the Huntington-Heberlein process may be
construed as applying to the Savelsberg. The charge is prepared
in the same way, the method of operating the converters is the
same, and the results of the reactions in the converters are the
same. The litigation which is pending between the two interests,
Messrs. Huntington and Heberlein claiming that Savelsberg in-
fringes their patents, will be, however, a deterrent to the extension
of the Savelsberg process until that matter be settled.

The Carmichael-Bradford process may be dismissed with a
few words. It is similar to the Savelsberg, except that gypsum
is used instead of limestone. It is somewhat more expensive
because the gypsum has to be ground and calcined. The process
works efficiently at Broken Hill, but it can hardly be of general
application, because gypsum is likely to be too expensive, except
in a few favored localities. The ability to utilize the converter
gases for the manufacture of sulphuric acid will cut no great
figure, save in exceptional cases, as at Broken Hill, and anyway
the gases of the other processes can be utilized for the same
purpose, which is in fact being done in connection with the
Huntington-Heberlein process in Silesia.

The cost of desulphurizing a ton of galena concentrate by the
Carmichael-Bradford process is estimated by the company con-
trolling the patents as follows, labor being reckoned at $1.80 per
eight hours, gypsum at $2.40 per 2240 lb., and coal at $8.40 per
2240 lb.:

0.25 ton of gypsum $0.60

Dehydrating and granulating gypsum 48

Drying mixture of ore and gypsum 12

210 LEAD SMELTING AND REFINING

Converting $0.24

Spalling sintered material 12

0.01 ton coal 08

Total $1 .64

The value of the lime in the sintered product is credited at
12c., making the net cost $1.52 per 2240 Ib. of ore.

The cost allowed for converting may be explained by the
more rapid action that appears to be attained with the ores of
Broken Hill than with some ores that are treated in North America,
but the low figure estimated for spalling the sintered material
appears to be highly doubtful.

The theory of the lime-roasting processes is not yet well
established. It is recognized that the explanation offered by
Huntington and Heberlein in their original patent specification
is erroneous. There is no good evidence in their process, or any
other, of the formation of the higher oxide of lime, which they

At the present time there are two views. In one, formulated
most explicitly by Professor Borchers, there is formed in this
process a plumbate of calcium, which is an active oxidizing agent.
A formation of this substance was also described by Carmichael
in his original patent, but he considered it to be the final product,
not the active oxidizing agent.

In the other view, the lime, or limestone, serves merely as a
diluent of the charge, enabling the air to obtain access to the
particles of galena, without liquefaction of the latter. The oxi-
dation of the lead sulphide is therefore effected chiefly by the
air, and the process is analogous to what takes place in the besse-
mer converter or in the Germot process of smelting, or perhaps
more closely to what might happen in an ordinary roasting
furnace, provided with a porous hearth, through which the air
supply would be introduced. Roasting furnaces of that design
have been proposed, and in fact such a construction is now
being tested for blende roasting in Kansas.

Up to the present time, the evidence is surely too incomplete
to enable a definite conclusion to be reached. Some facts may,
however, be stated.

There is clearly reaction to a certain extent between lead
sulphide and lead sulphate, as in the reverberatory smelting

LIME-ROASTING OF GALENA 211

furnace, because prills of metallic lead are to be observed in the
lime-roasted charge.

There is a formation of sulphuric acid in the lime-roasting,
upon the oxidizing effect of which Savelsberg lays considerable
stress, since its action is to be observed on the iron work in
which it condenses.

Calcium sulphate, which is present in all of the processes,
being specifically added in the Carmichael-Bradford, evidently
plays an important chemical part, because not only is the sulphur
trioxide expelled from the artificial gypsum, but also it is to a
certain extent expelled from the natural gypsum, which is added
in the Carmichael-Bradford process; in other words, more sulphur
is given off by the charge than is contained by the metallic sul-
phides alone.

Further evidence that lime does indeed play a chemical part
in the reaction is presented by the phenomena of lime-roasting
in clay dishes in the assay muffle, wherein the air is certainly
not blown through the charge, which is simply exposed to super-
ficial oxidation as in ordinary roasting.

The desulphurized charge dropped from the pot is certainly
at much below the temperature of fusion, even in the interior,
but we have no evidence of the precise temperature condition
during the process itself.

Pyrite and even zinc blende in the ore are completely oxidized.
This, at least, indicates intense atmospheric action.

The papers by Borchers,1 Doeltz,2 Guillemain,3 and Hutch-
ings 4 may profitably be studied in connection with the reactions
involved in lime-roasting. The conclusion will be, however, that
their precise nature has not yet been determined. In view of
the great interest that has been awakened by this new departure
in the metallurgy of lead, it is to be expected that much experi-
mental work will be devoted to it, which will throw light upon
its principles, and possibly develop it from a mere process of
desulphurization into one which will yield a final product in a
single operation.

1 Metallurgie, 1905,11, i, 1-6; Engineering and Mining Journal, Sept. 2, 1905.

2 Metallurgie, 1905, II, 19; Engineering and Mining Journal, Jan. 27, 1906.

3 Metallurgie, 1905; Sept. 22, 1905; Engineering and Mining Journal,
March 10, 1906.

4 Engineering and Mining Journal, Oct. 21, 1905.

PART VI
OTHER METHODS OF SMELTING

THE BORMETTES METHOD OF LEAD AND COPPER
SMELTING l

BY ALFREDO LOTTI

(September 30, 1905)

It is well known that, in order to obtain a proper fusion in
lead and copper ore-smelting, it is not only advantageous, but
often indispensable, that a suitable proportion of slag be added
to the charge. In the treatment of copper matte in the con-
verter, the total quantity of slag must be resmelted, inasmuch
as it always retains a notable quantity of the metal; while in
the smelting of lead ore in the blast furnace, the addition of slag
is mainly intended to facilitate the operation, avoiding the use
of strong air pressure and thus diminishing the loss of lead.
The proportion of slag required sometimes amounts to 30 to 35
per cent, of the weight of the ore.

Inasmuch as the slag is usually added in lump form, cold, its
original heat (about 400 calories per kilogram) is completely lost
and an intimate mixture with the charge cannot be obtained.
For this reason, I have studied the agglomeration of lead and
copper ores with fused slag, employing a variable proportion
according to the nature of the ore treated. In the majority of
cases, and with some slight modifications in each particular case,
by incorporating the dry or slightly moistened mineral with the
predetermined quantity of liquid slag, and by rapidly stirring
the mixture so as to secure a proper subdivision of the slag and
the mineral, there is produced a spongy material, largely com-
posed of small pieces, together with a simultaneous evolution of
dense fumes of sulphur, sulphur dioxide, and sulphur trioxide.
By submitting this spongy material to an air blast, the sulphur
of the mineral is burned, the temperature rising in the interior
of the mass to a clear red heat. Copious fumes of sulphur dioxide
and trioxide are given off, and at times a yellowish vapor of
sulphur, which condenses in drops, especially if the ore is pyritous.

i Translated by W. R. Ingalls.
215

216 LEAD SMELTING AND REFINING

At the end of from one to three hours, according to the quan-
tity of sulphur contained in the material under treatment and
the amount of the air pressure, the desulphurization of the ore,
so far as it has come in contact with the air, is completed, and
the mass, now thoroughly agglomerated, forms a spongy but
compact block. It is then only necessary to break it up and
smelt it with the requisite quantity of flux and coke. The
physical condition of the material is conducive to a rapid and
economical smelting, while the mixture of the sulphide, sulphate
and oxide leads to a favorable reaction in the furnace.

In employing this method, it sometimes happens that ores
rich in sulphur produce during the smelting a little more matte
than when the ordinary system of roasting is employed. In such
instances, in order to avoid or to diminish the cost of re-treatment
of the matte, it is best to agglomerate a portion thereof with the
crude mineral and the slag. This has the advantage of oxidizing
the matte, which acts as a ferruginous flux in the smelting.

The system described above leads to considerable economy,
especially in roasting, as the heat of the scoria, together with
that given off in the combustion of the sulphur, is almost always
sufficient for the agglomeration and desulphurization of the
mineral; while, moreover, it reduces the cost of smelting in the
blast furnace. Although the primary desulphurization is only
partial (about 50 per cent.), it continues in the blast furnace, since
the mineral, agglomerated with the slag, assumes a spongy form
and thereby presents an increased surface to the action of the
air. The sulphur also acts as a fuel and does not produce an
excessive quantity of matte.

The system will prove especially useful in the treatment of
argentiferous lead ore, since, by avoiding the calcination in a
reverberatory furnace, loss of silver is diminished. It appears,
however, that, contrary to the reactions which occur in the
Huntington-Heberlein process, a calcareous or basic gangue is
not favorable to this process, if the proportion be too great.

The following comparison has been made in the case of an
ore containing 62 to 65 per cent, of lead, 16 to 17 per cent, sul-
phur, 10 to 11 per cent, zinc, 0.4 per cent, copper, and 0.222 per
cent, silver, in which connection it is to be remarked that, in
general, the less zinc there is in the ore the better are the re-
sults.

OTHER METHODS OF SMELTING

217

FIG. 21. — Elevation and Plan of Converting Chambers.

218 LEAD SMELTING AND REFINING

Ordinary Method. — Roast-reduction. Cost per 1000 kg. of
crude ore:

1. Roasting in reverberatory furnace:

Labor $0.70

Fuel 1.50

Repairs and supplies 05

- $2.25

2. Smelting in water-jacket:

Labor $1.01

Fuel 2.20

Repairs and supplies 03

Fluxes 50

3.74

Total $5.99

Bormettes Method. — Agglomeration with slag, pneumatic de-
sulphurization and smelting in water-jacket:

1. Agglomeration and desulphurization:

Labor $0.42

Repairs and supplies 0.05

- $0.47

2. Smelting in water-jacket:

Labor $0.90

Fuel 1.91

Repairs and supplies 03

Fluxes 42

3.26

Total $3.73

This shows a difference in favor of the new method of $2.26
per ton of ore, without taking into account the savings realized
by a much more speedy handling of the operation, which would
further reduce the cost to approximately $2.50 per ton.

In the above figures, no account has been taken of general
expenses, which per ton of ore are reduced because of the greater
rapidity of the process, enabling a larger quantity of ore to be
smelted in a given time. Making allowance for this, the saving
will amount to an average of $2.40 per 1000 kg., a figure which
will naturally vary according to the prices for fuel, labor, and
the quantity of matte which it may be necessary to re-treat.

OTHER METHODS OF SMELTING

219

FIG. 22. — Details of Transfer Cars.

220

LEAD SMELTING AND REFININQ

If the quantity of matte does not exceed 10 per cent, of the
weight of the ore, it can be desulphurized by admixture with the
ore, without use of other fuel. If, however, the proportion of
matte rises to 20 parts per 100 parts of ore (a maximum which
ought not to be reached in good working), it is necessary to
roast a portion of it. Under unfavorable conditions, consequently,
the saving effected by this process may be reduced to $2 @ $2.20

FIG. 23. — Latest Form of Converter. (Section on A B.)

per 1000 kg., and even to as little as $1.40 @ $1.60. The above
reckonings are, however, without taking any account of the
higher extraction of lead and silver, which is one of the great
advantages of the Bormettes process.

The technical results obtained in the smelting of an ore of
the above mentioned composition are as follows:

OTHER METHODS OF SMELTING

221

; .

ORDINARY
METHOD

BORMETTES

METHOD

Coke per cent of the charge

Hlust pressure water gage • .

12 to 20 cm

12 to 14 cm

Tons of charge smelted per 24 hr
Tons of ore smelted per 24 hr

25
10

Lead assay of slag

0 80 to 0 90%

0.20 to 0 40%

Matte-fall, per cent, of ore charged

5 to 10

10 to 15

Lead extraction

90%

Silver extraction

95%

98^

. •

FIG. 24. — Latest Form of Converter. (Section on C D.)

The higher extractions of lead and silver are explained by the
fact that the loss of metals in roasting is reduced, while, more-
over, the slags from the blast furnace are poorer than in the
ordinary process of smelting. The economy in coke results from
the greater quantity of sulphur which is utilized as fuel, and
from the increased fusibility of the charge for the blast furnace.

222

LEAD SMELTING AND REFINING

The new system of desulphurization enables the charge to be
smelted with a less quantity of fresh flux, by the employment in
its place of a greater proportion of foul slag. The reduction in
the necessary amount of flux is due not only to the increased
fusibility of the agglomerated charge, but principally to the fact
that in this system the formation of silicates of lead (which are
produced abundantly in ordinary slag-roasting) is almost nil. It
is therefore unnecessary to employ basic fluxes in order to reduce
scorified lead.

FIG. 25. — Latest Form of Converter. (Plan.)

The losses of metal in the desulphurization are less than in
the ordinary method, because the crude mineral remains only a
short time (from one to three hours) in the apparatus for desul-
phurization and agglomeration, and the temperature of the
process is lower. The blast-furnace slags are poorer, because1
there is no formation of silicate of lead during the agglomeration.

The Bormettes method, in so far as the treatment of lead ore
is concerned, may be considered a combination process of roast-
reaction, of roast-reduction, and of precipitation-smelting. It is

OTHER METHODS OF SMELTING 223

not, however, restricted to the treatment of lead ore. It may
also be applied to the smelting of pyritous copper-bearing ores.
In an experiment with cupriferous pyrites, containing 20 to 25
per cent, sulphur, it succeeded in agglomerating and smelting
them without use of any fuel for calcination, effecting a perfect
smelting, analogous to pyrite smelting, with the production of a
matte of sufficient degree of concentration.

The first cost of plant installation is very much reduced by
the Bormettes method, inasmuch as the ordinary roasting fur-
naces are almost entirely dispensed with, apparatus being sub-
stituted for them which cost only one-third or one-fourth as
much as ordinary furnaces. The process presents the advantage,
moreover, of being put into immediate operation, without any
expenditure of excess fuel.

The apparatus required in the process is illustrated in Figs.
21-25. The apparatus for desulphurization and agglomeration
consists of a cast-iron box, composed of four vertical walls, of
which two incline slightly toward the front. These inclined
walls carry the air-boxes. The other two walls are formed, the
one in front by the doors which give access to the interior, and
the other in the rear by a straight plate. The whole arrangement
is surmounted by a hood. The four pieces when assembled form
a box without bottom. Several of these boxes are combined as
a battery. The pots in which the agglomeration and desulphuri-
zation are effected are moved into these boxes on suitable cars,
in the manner shown in the first engraving. A later and more
improved form is shown, however, in Figs. 23-25.

This process, which is the invention of A. Lotti and has been
patented in all the principal countries, is in successful use at the
works of the Socie*te Anonyme des Mines de Bormettes, at Bor-
mettes, La Londe (Var), France. Negotiations are now in
progress with respect to its introduction elsewhere in Europe.

THE GERMOT PROCESS1

BY WALTER RENTON INGALLS

(November 1, 1902)

According to F. Laur, in the Echo des Mines (these notes are
abstracted from Oest. Zeit., L., xl, 55, October 4, 1902), A. Germot,
of Clichy, France, made experiments some years ago upon the
production of white lead directly from galena. These led Catelin
to attempt the recovery of metallic lead in a similar way. If
air be blown in proper quantity into a fused mass of lead sulphide
the following reaction takes place:

2PbS + 2O = SO2 + Pb + PbS.

Thus one-half of the lead is reduced, and it is found collects
all the silver of the ore; the other half is sublimed as lead sul-
phide, which is free from silver. The reaction is exothermic to
the extent that the burning of one-half the sulphur of a charge
should theoretically develop sufficient heat to volatilize half of
the charge and smelt the other half. This is almost done in
practice with very rich galena, but not so with poorer ore. The
temperature of the furnace must be maintained at about 1100
deg. C. throughout the whole operation, and there are the usual
losses of heat by radiation, absorption by the nitrogen of the air,
etc. Deficiencies in heat are supplied by burning some of the
ore to white lead, which is mixed with the black fume (PbS) and
by the well-known reactions reduced to metal with evolution of
sulphur dioxide. The final result is therefore the production of
(1) pig lead enriched in silver; (2) pig lead free from silver; (3) a
leady slag; and (4) sulphur dioxide. In the case of ores contain-
ing less than 75 per cent. Pb the gangue forms first a little skin
and then a thick hard crust which soon interferes with the opera-
tion, especially if the ore be zinkiferous. This difficulty is over-

1 As originally published the title of this article was " Lead-Smelting
without Fuel." In this connection reference may well be made to Hannay's
experiments and theories, Transactions Institution of Mining and Metallurgy,
II, 188, and Huntington's discussion, ibid., p. 217.

224

OTHER METHODS OF SMELTING

225

come by increasing the temperature or by fluxing the ore so as
to produce a fusible slag. A leady slag is always easily produced;
this is the only by-product of the process. The theoretical reac-
tion requires 600 cu. m. of air, assuming a delivery of 50 per cent,
from the blower, and at one atmosphere pressure involves the
expenditure of 18 h.p. per 1000 kg. of galena per hour.

FIG. 26. — Plan and Elevation of Smelting Plant at Clichy.

The arrangement of the plant at Clichy is shown diagrammati-
cally in Fig. 26. There is a round shaft furnace, 0.54 meter in
diameter and 4.5 meters high. Power is supplied to the blower
C through the pulley G and the shaft DD. The compressed air
is accumulated in the reservoir R, whence it is conducted by
the pipe to the tuyere which is suspended inside of the furnace
by means of a chain, whereby it can be raised or lowered. Ot

226 LEAD SMELTING AND REFINING

and O2 are tap-holes. L is a door and N an observation tube. A
is the charge tube. X is the pipe which conveys the gas and
fume to the condensation chambers. T is the pipe through
which the waste gases are drawn. V is the exhauster and S is
the chimney. Kj and K2 are tilting crucible furnaces for melt-
ing lead and galena.

After the furnace has been properly heated, 100 kg. of lead
melted in Kt are poured in through the cast-iron pipe P, and
after that about 200 kg. of pure, thoroughly melted galena from
K2. Ore containing 70 to 80 per cent. Pb must be used for this
purpose. The blast of air is then introduced into the molten
galena, and from 1000 to 3000 kg. of ore is gradually charged in
through the tube A. During this operation black fume (PbS)
collects in the condensation chamber. All outlets are closed
against the external air. If the air blast is properly adjusted,
nothing but black fume is produced; if it begins to become light
colored, charging is discontinued and the blast of air is shut off.
Lead is then tapped through O2, which is about 0.2 meter above
the hearth, so there is always a bath of lead in the bottom of the
furnace; but it is advisable now and then to tap off some through
Oj, so as gradually to heat up the bottom of the furnace. Hearth
accretions are also removed through Or The lead is tapped off
through O2 until matte appears. The tap hole is then closed,
the tuyere is lowered and the blast is turned into the lead in order
to oxidize it and completely desulphurize the sulphur combina-
tions, which is quickly done. The oxide of lead is scorified as a
very fusible slag, which is tapped off through O2, and more ore
is then charged in upon the lead bath and the cycle of operations
is begun again.

PART VII

DUST AND FUME RECOVERY
FLUES, CHAMBERS AND BAG-HOUSES

DUST CHAMBER DESIGN

BY MAX J. WELCH

(September 1, 1904)

Only a few years ago smelting companies began to recognize
the advantage of large chambers for collecting flue dust and
condensing fumes. The object is threefold: First, profit; second,
to prevent law suits with surrounding agricultural interests;
third, cleanliness about the plant. It is my object at present to
discuss the materials used in construction and general types of
cross-section.

Most of the old types of chambers are built after one general
pattern, namely, brick or stone side walls and arch roof, with
iron buckstays and tie rods. The above type is now nearly out
of use, because it is short-lived, expensive, and dangerous to
repair, while the steel and masonry are not used to good advan-
tage in strength of cross-section.

With the introduction of concrete and expanded metal began
a new era of dust-chamber construction. It was found that a
skeleton of steel with cement plaster is very strong, light and
cheap. The first flue of the type shown in Fig. 29 was built after
the design of E. H. Messiter, at the Arkansas Valley smelter in
Colorado. This flue was in commission several years, conveying
sulphurous gases from the reverberatory roaster plant. The
same company decided, in 1900, to enlarge and entirely rebuild
its dust-chamber system, and three types of cross-section were
adopted to meet the various conditions. All three types were of
cement and steel construction.

The first type, shown in Fig. 27, is placed directly behind the
blast furnaces. The cross-section is 273 sq. ft. area, being de-
signed for a 10-furnace lead smelter. The back part is formed
upon the slope of the hillside and paved with 2.5 in. of brick.
The front part is of ribbed cast-iron plates. Ninety per cent, of
the flue dust is collected in this chamber and is removed, through
sliding doors, into tram cars. There is a little knack in designing

229

230

LEAD SMELTING AND REFINING

a door to retain flue dust. It is simply to make the bottom sill
of the door frame horizontal for a space of about 1 in. outside of
the door slide.

The front part of the chamber, Fig. 27, is of expanded metal
and cement. The top -is of 20-in. I-beams, spanning a distance
of 24 ft. with 15-in. cross-beams and 3 in. of concrete floor resting
upon the bottom flanges of the beams. This heavy construction
forms the foundation for the charging floor, bins, scales, etc.

While dwelling upon this type of construction I wish to men-
tion a most important point, that of the proper factor of safety.

FIG. 27. — Rectangular form of Concrete Dust Chamber.

Flue dust, collected near the blast furnace, weighs from 80 to
100 Ib. per cubic foot, and the steel supports should be designed
for 16,000 Ib. extreme fiber stress, when the chamber is three-
quarters full of dust. If the dust is allowed to accumulate
beyond this point, the steel, being well designed, should not be
overstrained. Discussions as to strains in bins have been aired
by the engineering profession, but the present question is "Where
is a dust chamber a bin?" Experience shows that bin construc-
tion should be adopted behind, or in close proximity to, the blast
furnaces.

DUST AND FUME RECOVERY

231

Fig. 28 shows the second type of hopper-bottom flue adopted.
It is of very light construction, of 274 sq. ft. area in the clear.
The beginning of this flue being 473 ft. from the blast furnaces
removes all possibility of any material floor-load, as the dust is
light in weight and does not collect in large quantities. The
hopper-bottom floor is formed of 4-in. concrete slabs, in panels,
placed between 4-in. I-beams. Cast-iron door frames, with open-
ings 12 x 16 in., are placed on 5-ft. centers. The concrete floor
is tamped in place around the frames. The side walls and roof
are built of 1 -in. angles, expanded metal, and plastered to 2.5 in.

FIG. 28. — Arched form of Concrete Dust Chamber.

thickness. At every 10-ft. distance, pilaster ribs built of 2-in.
angles, latticed and plastered, form the wind-bracing and arch
roof support.

Fig. 29 shows the beehive construction. This chamber is of
253 sq. ft. cross-sectional area. It is built of 2-in. channels, placed
16 in. centers, tied with 1 x 0.125 in. steel strips. The object
of the strips is to support the 2-in. channels during erection.
No. 27 gage expanded metal lath was wired to the inside of the
channels and the whole plastered to a thickness of 3 in. The
inside coat was plastered first with portland cement and sand,

232 LEAD SMELTING AND REFINING

one to three, with about 5 per cent. lime. The filling between
ribs is one to four, and the outside coat one to three.

The above types of dust chamber have been in use over three
years at Leadville. Cement and concrete, in conjunction with
steel, have been used in Utah, Montana and Arizona, in various
types of cross-section. The results show clearly where not to
use cement; namely, where condensing sulphur fumes come in
contact with the walls, or where moisture collects, forming sul-
phuric acid. The reason is that portland cement and lime
mortar contain calcium hydrate, which takes up sulphur from
the fumes, forming calcium sulphate. In condensing chambers,
this calcium sulphate takes up water, forming gypsum, which
expands and peels off.

FIG. 29. — Beehive form of Concrete Dust Chamber.

In materials of construction it is rather difficult to get some-
thing that will stand the action of sulphur fumes perfectly. The
lime mortar joints in the old types of brick flues are soon eaten
away. The arches become weak and fall down. I noted a sheet
steel condensing system, where in one year the No. 12 steel was
nearly eaten through. With a view of profiting by past expe-
rience, let us consider the acid-proof materials of construction,
namely, brick, adobe mortar, fire-clay, and acid-proof paint.
Also, let us consider at what place in a dust-chamber system

DUST AND FUME RECOVERY 233

are we to take the proper precaution in the use of these mate-
rials.

At smelting plants, both copper and lead, it is found that
near the blast furnaces the gases remain hot and dry, so that
concrete, brick or stone, or steel, can safely be used. Lead-blast
furnace gases will not injure such construction at a distance of
6 or 8 ft. away from the furnaces. For copper furnaces, roasters
or pyritic smelting, concrete or lime mortar construction should
be limited to within 200 or 300 ft. of the furnaces.

Another type of settling chamber is 20 ft. square in the
clear, with concrete floor between beams and steel hopper bottom.
This chamber is built within 150 ft. distance from the blast fur-
naces, and is one of the types used at the Shannon Copper Com-
pany's plant at Clifton, Arizona. After passing the 200-ft. mark,
there is no need of expensive hopper design. The amount of
flue dust settled beyond this point is so small that it is a better
investment to provide only small side doors through which the
dust can be removed. The ideal arrangement is to have a system
of condensing chambers, so separated by dampers that either set
can be thrown out for a short time for cleaning purposes, and
the whole system can be thrown in for best efficiency.

As to cross-section for condensing chambers, I consider that
the following will come near to meeting the requirements. One,
four, and six, concrete foundation; tile drainage; 9-in. brick walls,
laid in adobe mortar, pointed on the outside with lime mortar;
occasional strips of expanded metal flooring laid in joints; the
necessary pilasters to take care of the size of cross-section adopted;
the top covered with unpainted corrugated iron, over which is
tamped a concrete roof, nearly flat; concrete to contain corru-
gated bars in accordance with light floor construction; and lastly,
the corrugated iron to have two coats of graphite paint on under
side.

The above type of roof is used under slightly different condi-
tions over the immense dust chamber of the new Copper Queen
smelter at Douglas, Arizona. The paint is an important consid-
eration. Steel work imbedded in concrete should never be
painted, but all steel exposed to fumes should be covered by
graphite paint. Tests made by the United States Graphite Com-
pany show that for stack work the paint, when exposed to acid
gases, under as high a temperature as 700 deg. F., will wear well.

CONCRETE IN METALLURGICAL CONSTRUCTION1

BY HENRY W. EDWARDS

The construction of concrete flues of the section shown in
Fig. 31 gives better results than that shown in Fig. 30, being less
liable to collapse. It costs somewhat more to build owing to
the greater complication of the crib, which, in both cases, consists
of an interior core only. For work 4 in. in thickness and under,
I recommend the use of rock or slag crushed to pass through a
1.5-in. ring. Although concrete is not very refractory, it will
easily withstand the heat of the gases from a set of ordinary

FIGS. 30 and 31. — Sections of Concrete Flues.

lead- or copper-smelting blast furnaces, or from a battery of
calcining or roasting furnaces. I have never noticed that it is
attacked in any way by sulphur dioxide or other furnace gas.

Shapes the most complicated to suit all tastes in dust chambers
can be constructed of concrete. The least suitable design, so
far as the construction itself is concerned, is a long, wide, straight-
walled, empty chamber, which is apt to collapse, either inwards
or outwards, and, although the outward movement can be

1 Excerpt from a paper, "Concrete in Mining and Metallurgical Engineer-
ing," Transactions American Institute of Mining Engineers, XXXV (1905),
p. 60.

234

DUST AND FUME RECOVERY

235

prevented by a system of light buckstays and tie-rods, the ten-
dency to collapse inwards is not so simply controlled in the
absence of transverse baffle walls. The tendency, so far as the
collection of mechanical flue dust is concerned, appears to be
towards a large empty chamber, without baffles, etc., in which
the velocity of the air currents is reduced to a minimum, and
the dust allowed to settle. In the absence of transverse baffle
walls to counteract the collapsing tendency, it seems best to
design the chamber with a number of stout concrete columns at
suitable intervals along the side and end walls — the walls them-
selves being made only a few inches thick with woven-wire screen
or "expanded metal" buried within them. The wire skeleton
should also be embedded into the columns in order to prevent

FIG. 32. — Concrete Dust Chamber at the Guillermo Smelting Works, Palo-
mares, Spain. (Horizontal section.)

the separation of wall and the columns. This method of con-
structing is one that I have followed with very satisfactory
results as far as the construction itself is concerned.

Figs. 32 and 33 show a chamber designed and erected at the
Don Guillermo Smelting Works at Palomares, Province of Mureia,
Spain. Figs. 34 and 35 show a design for the smelter at Murray
Mine, Sudbury, Ontario, in which the columns are hollow, thus
economizing concrete material. For work of this kind the col-
umns are built first and the wire netting stretched from column
to column and partly buried within them. The crib is then built
on each side of the netting, a gang of men working from both
sides, and is built up a yard or so at a time as the work progresses.
Doors of good size should be provided for entrance into the

236

LEAD SMELTING AND REFINING

chamber, and as they will seldom be opened there is no need for
expensive fastenings or hinges.

Foundations for Dynamos and other Electrical Machinery. —

4)* Thick

FIG. 33. — Concrete Dust Cham-
ber at the Guillermo Smelting
Works, Palomares, Spain. (End
elevation.)

Dry concrete is a poor conductor of electricity, but when wet it
becomes a fairly good conductor. Therefore, if it be necessary
to insulate the electrical apparatus, the concrete should be covered
with a layer of asphalt.

Fie. 34. — Concrete Dust Chamber designed for smelter at Murray Mine,
Sudbuiy, Ontario, Can. There are eight 9-ft. sections in the plan.

DUST AND FUME RECOVERY

237

Chimney Bases. — Fig. 36 shows the base for the 90-ft. brick-
stack at Don Guillermo. The resemblance to masonry is given
by nailing strips of wood on the inside of the crib.

Retaining-Walls. — Figs. 37, 38, and 39 show three different
styles of retaining-walls, according to location. These walls are
shown in section only, and show the placing of the iron reen-
forcements. Retaining-walls are best built in panels (each panel
being a day's work), for the reason that horizontal joints in the
concrete are thereby avoided. The alternate panels should be
built first and the intermediate spaces filled in afterward. Should
there be water behind the wall it is best to insert a few small

FIG. 35. — Concrete Dust Chamber designed for smelter at Murray Mine,
Sudbury, Ontario, Can. (End elevation.)

pipes through the wall, in order to carry it off; this precaution is
particularly important in places where the natural surface of the
ground meets the wall, as shown in Figs. 37 and 38. If a wooden
building is to be erected on the retaining- wall, it is best to bury
a few 0.75-in. bolts vertically in the top of the wall, by which a
wooden coping may be secured (see Figs. 37, 38, and 39), which
forms a good commencement for the carpenter work.

Minimum thickness for a retaining-wall, having a liberal
quantity of iron embedded therein, is 20 in. at the bottom and
10 in. at the top, with the taper preferably on the inner face.
In the absence of interior strengthening irons the thickness of

238

LEAD SMELTING AND REFINING

the wall at the bottom should never be less than one-fourth the
total hight, and at the top one-seventh of the hight; unless
very liberal iron bracing be used, the dimensions can hardly be
reduced to less than one-seventh and one-tenth respectively.
Unbraced retaining-walls are more stable with the batter on the
outer face. Dry clay is the most treacherous material that can

FIQ. 36. — Concrete Base for a 90-ft. Chimney at the Guillermo Smelting
Works, Palomares, Spain.

be had behind a retaining- wall, especially if it be beaten in, for
the reason that it is so prone to absorb moisture and swell, causing
an enormous side thrust against the wall. When this material
is to be retained it is best to build the wall superabundantly
strong — a precaution which applies even to a dry climate.

DUST AND FUME RECOVERY

239

because the bursting of a water-pipe may cause the damage.
In order to avoid horizontal joints it is best, wherever practicable,
to build the crib- work in its entirety before starting the concrete.
In a retaining-wall 3 ft. thick by 16 ft. high this is not practi-
cable. The supporting posts and struts can, however, be com-
pleted and the boards laid in as the wall grows, in order not to
interrupt the regular progress of the tamping. A good finish
may be produced on the exposed face of the wall by a few strokes
of the shovel up and down with its back against the crib.

»»»d Floor aPlneCoplaft

FIGS. 37, 38, and 39. — Retaining-Walls of Concrete.

In conclusion I wish to state that this paper is not written
for the instruction of the civil engineer, or for those who have
special experience in this line; but rather for the mining engineer
or metallurgist whose training is not very deep in this direction,
and who is so often thrown upon his own resources in the wil-
derness, and who might be glad of a few practical suggestions
from one who has been in a like predicament.

CONCRETE FLUES1

BY EDWIN H. MESSITER

(September, 1904)

Under the heading " Flues," Mr. Edwards refers to the Bee-
hive construction, a cross-section of which is shown in Fig. 31
of his paper. A flue similar to this was designed by me about
six years ago,2 and in which the walls, though much thinner than
those described by Mr. Edwards, gave entire satisfaction. These
walls, from 2.25 in. thick throughout in the smaller flues to
3.25 in. in the larger, were built by plastering the cement mortar
on expanded-metal lath, without the use of any forms or cribs
whatever, at a cost of labor generally less than $1 per sq. yd. of
wall. Of course, where plasterers cannot be obtained on reason-
able terms, the cement can be molded between wooden forms,,
though it is difficult to see how it can be done with an interior
core only, as stated by Mr. Edwards.

In regard to the effect of sulphur dioxide and furnace gases
on the cement, I have found that in certain cases this is a matter
which must be given very careful attention. Where there is.
sufficient heat to prevent the existence of condensed moisture
inside of the flue, there is apparently no action whatever on the
cement, but if the concrete is wet, it is rapidly rotted by these
gases. At points near the furnaces there is generally sufficient
heat not only to prevent internal condensation of the aqueous
vapor always present in the gases, but also to evaporate water
from rain or snow falling on the outside of the flue. Further
along a point is reached where rain-water will percolate through
minute cracks caused by expansion and contraction, and reach
the interior even though internal condensation does not occur
there in dry weather. From this point to the end of the flue the

1 A Discussion of the Paper by Henry W. Edwards, on " Concrete in
Mining and Metallurgical Engineering," Transactions of the American Institute
of Mining Engineers, XXXV.

1 Engineering News, Nov 30, 1899, and U. S. Patent No. 665,250, Jan. 1
1901.

240

DUST AND FUME RECOVERY 241

roof must be coated on the outside with asphalt paint or other
impervious material. In very long flues a point may be reached
where moisture will condense on the inside of the walls in cold
weather. From this point to the end of the flue it is essential
to protect the interior with an acid-resisting paint, of which two
or more coats will be necessary. For the first coat a material
containing little or no linseed oil is best, as I am informed that
the lime in the cement attacks the oil. For this purpose I have
used ebonite varnish, and for the succeeding coats durable
metal-coating. The first coat will require about 1 gal. of material
for each 100 sq. ft. of surface.

In one of the earliest long flues built of cement in this country,
a small part near the chimney was damaged as a result of failure
to apply the protective coating, the necessity for it not having
been recognized at the time of its construction. It may be said,
in passing, that other long brick flues built prior to that time
were just as badly attacked at points remote from the furnaces.
In order to reduce the amount of flue subject to condensation,
the plastered flues have been built with double lath having an
intervening air-space in the middle of the wall.

In building thin walls of cement, such as flue walls, it is
particularly important to prevent them from drying before the
cement has combined with all the water it needs. For this
reason the work should be sprinkled freely until the cement is
fully set. Much work of this class has been ruined through
ignorance by fires built near the walls in cold weather, which
caused the mortar to shell off in a short time.

The great saving in cost of construction, which the concrete-
steel flue makes possible, will doubtless cause it to supersede
other types to even a greater extent than it has already done.
If properly designed this type of construction reduces the cost
of flues by about one-half. Moreover, the concrete-steel flue is
a tight flue as compared with one built of brick. There is a
serious leakage through the walls of the brick flues which is not
easily observed in flues under suction as most flues are, but
when a brick flue is under pressure from a fan the leakage is
surprisingly apparent. In flues operating by chimney-draft the
entrance of cold air must cause a considerable loss in the efficiency
of the chimney, a disadvantage which would largely be obviated
by the use of the concrete-steel flue.

CONCRETE FLUES1

BY FRANCIS T. HAVAKD

In discussion of Mr. Edwards's interesting and valuable paper,
I beg to submit the following notes concerning the advantages
and disadvantages of the concrete flues and stacks at the plant
of the Anhaltische Blei- und Silber-werke. The flues and smaller
stacks at the works were constructed of concrete consisting
generally of one part of cement to seven parts of sand and jig-
tailings but, in the case of the under-mentioned metal concrete
slabs, of one part of cement to four parts of sand and tailings.
The cost of constructing the concrete flue approximated 5 marks
per sq. m. of area (equivalent to $0.11 per sq. ft.).

Effect of Heat. — A temperature above 100 deg. C. caused the
concrete to crack destructively. Neutral furnace gases at 120
deg. C., passing through an independent concrete flue and stack,
caused so much damage by the formation of cracks that, after
two years of use, the stack, constructed of pipes 4 in. thick,
required thorough repairing and auxiliary ties for every foot of
hight.

Effect of Flue Gases and Moisture. — The sides of the main
flue, made of blocks of 6-in. hollow wall-sections, 100 cm. by
50 cm. in area, were covered with 2-in. or 1-in. slabs of metal
concrete. In cases where the flue was protected on the outside
by a wooden or tiled roof, and inside by an acid-proof paint,
consisting of water-glass and asbestos, the concrete has not been
appreciably affected. In another case, where the protective cover,
both inside and outside, was of asphalt only, the concrete was
badly corroded and cracked at the end of three years. In a
third case, in which the concrete was unprotected from both
atmospheric influence on the outside, and furnace gases on the
inside, the flue was quite destroyed at the end of three years.

1 A discussion of the paper of Henry W. Edwards, on "Concrete in
Mining and Metallurgical Engineering," Transactions of the American Institute
of Mining Engineers, XXXV.

242

DUST AND FUME RECOVERY 243

That portion of the protected concrete flue, near the main stack,
which came in contact only with dry, cold gases was not affected
at all.

Gases alone, such as sulphur dioxide, sulphur trioxide, and
others, do not affect concrete; neither is the usual quantity of
moisture in furnace gases sufficient to damage concrete; but
should moisture penetrate from the outside of the flue, and,
meeting gaseous SO2 or S03, form hydrous acids, then the concrete
will be corroded.

Effect of the Atmosphere Alone. — For outside construction
work, foundations and other structures not exposed to heat,
moist acid gases and chemicals, the concrete has maintained its
reputation for cheapness and durability.

Effect of Crystallization of Contained Salts. — In chemical
works, floors constructed of concrete are sometimes unsatisfac-
tory, for the reason that soluble salts, noticeably zinc sulphate,
will penetrate into the floor and, by crystallizing hi narrow
confines, cause the concrete to crack and the floor to rise in
places.

BAG-HOUSES FOR SAVING FUME

BY WALTER RENTON INGALLS

(July 15, 1905)

One of the most efficient methods of saving fume and very
fine dust in metallurgical practice is by filtration through cloth.
This idea is by no means a new one, having been proposed by
Dr. Percy, in his treatise on lead, page 449, but he makes no
mention of any attempt to apply it. Its first practical applica-
tion was found in the manufacture of zinc oxide direct from ores,
initially tried by Richard and Samuel T. Jones in 1850, and in
1851 modified by Samuel Wetherill into the process which con-
tinues in use at the present time in about the same form as origi-
nally. In 1878 a similar process for the manufacture of white
lead direct from galena was introduced at Joplin, Mo., by G. T.
Lewis and Eyre O. Bartlett, the latter of whom had previously
been engaged in the manufacture of zinc oxide in the East, from
which he obtained his idea of the similar manufacture of white
lead. The difference in the character of the ore and other con-
ditions, however, made it necessary to introduce numerous
modifications before the process became successful. The eventual
success of the process led to its application for filtration of the
fume from the blast furnaces at the works of the Globe Smelting
and Refining Company, at Denver, Colo., and later on for the
filtration of the fume from the Scotch hearths employed for the
smelting of galena in the vicinity of St. Louis.

In connection with the smelting of high-grade galena in
Scotch hearths, the bag-house is now a standard accessory. It
has received also considerable application in connection with
silver-lead blast-furnace smelting and in the desilverizing refin-
eries. Its field of usefulness is limited only by the character of
the gas to be filtered, it being a prerequisite that the gas contain
no constituent that will quickly destroy the fabric of which the
bags are made. Bags are also employed successfully for the
collection of dust in cyanide mills, and other works in which

244

DUST AND FUME RECOVERY

245

246 LEAD SMELTING AND REFINING

fine crushing is practised, for example, in the magnetic separating
works of the New Jersey Zinc Company, Franklin, N. J., where the
outlets of the Edison driers, through which the ore is passed, com-
municate with bag-filtering machines, in which the bags are caused
to revolve for the purpose of mechanical discharge. The filtration
of such dust is more troublesome than the filtration of furnace
fume, because the condensation of moisture causes the bags to
become soggy.

The standard bag-house employed in connection with furnace
work is a large room, in which the bags hang vertically, being
suspended from the top. The bags are simply tubes of cotton
or woolen (flannel) cloth, from 18 to 20 in. in diameter, and 20
to 35 ft. in length, most commonly about 30 ft. In the manu-
facture of zinc oxide, the fume-laden gas is conducted into the
house through sheet-iron pipes, with suitably arranged branches,
from nipples on which the bags are suspended, the lower end
of the bag being simply tied up until it is necessary to discharge
the filtered fume by shaking. In the bag-houses employed in
the metallurgy of lead, the fume is introduced at the bottom
into brick chambers, which are covered with sheet-iron plates,
provided with the necessary nipples; or else into hopper-bottom,
sheet-iron flues, with the necessary nipples on top. In either
case the bags are tied to the nipples, and are tied up tight at the
top, where they are suspended. When the fume is dislodged by
shaking the bags, it falls into the chamber or hopper at the
bottom, whence it is periodically removed.

The cost of attending a bag-house, collecting the fume, etc.,
varies from about lOc. per ton of ore smelted in a large plant like
the Globe, to about 25c. per ton in a Scotch-hearth plant treating
25 tons of ore per 24 hours.

No definite rules for the proportioning of filtering area to the
quantity of ore treated have been formulated. The correct
proportion must necessarily vary according to the volume of
gaseous products developed in the smelting of a ton of ore, the
percentage of dust and fume contained, and the frequency with
which the bags are shaken. It would appear, however, that in
blast furnaces and Scotch-hearth smelting a ratio of 1000 sq. ft.
per ton of ore would be sufficient under ordinary conditions.
The bag-house originally constructed at the Globe works had
about 250 sq. ft. of filtering area per ton of charge smelted, but

DUST AND FUME RECOVERY 247

this was subsequently increased, and Dr. lies, in his treatise on
lead-smelting, recommends an equipment which would correspond
to about 750 sq. ft. per ton of charge. At the Omaha works,
where the Brown-De Camp system was used, there was 80,000
sq. ft. of cloth for 10 furnaces 42 x 120 in., according to Hof-
man's " Metallurgy of Lead," which would give about 1000 sq. ft.
per ton of charge smelted, assuming an average of eight furnaces
to be in blast. A bag-house in a Scotch-hearth smeltery, at
St. Louis, had approximately 900 sq. ft. per ton of ore smelted.
At the Lone Elm works, at Joplin, the ratio was about 3500 sq. ft.
per ton of ore smelted, when the works were run at their maxi-
mum capacity. In the manufacture of zinc oxide the bag area
used to be from 150 to 200 sq. ft. per square foot of grate on
which the ore is burned, but at Palmerton, Pa. (the most modern
plant), the ratio is only 100 : 1. This corresponds to about 1400
sq. ft. of bag area per 2000 Ib. of charge worked on the grate.
In the manufacture of zinc-lead white at Canon City, Colo., the
ratio between bag area and grate area is 150 : 1.

Assuming the gas to be free, or nearly free, from sulphurous
fumes, the bags are made of unbleached muslin, varying in weight
from 0.4 to 0.7 oz. avoirdupois per square foot. The cloth should
have 42 to 48 threads per linear inch in the warp and the same
number in the woof. A kind of cloth commonly used in good
practice weighs 0.6 oz. per square foot and has 46 threads per
linear inch in both the warp and the woof.

The bags should be 18 to 20 in. in diameter. Therefore the
cloth should be of such width as to make that diameter with
only one seam, allowing for the lap. Cloth 62 in. in width is
most convenient. It costs 4 to 5c. per yard. The seam is
made by lapping the edges about 1 in., or by turning over the
edges and then lapping, in the latter case the stitches passing
through four thicknesses of the cloth. It should be sewed
with No. 50 linen thread, making two rows of double lock-
stitches.

The thimbles to which the bags are fastened should be of
No. 10 sheet steel, the rim being formed by turning over a ring
of 0.25 in. wire. The bags are tied on with 2-in. strips of
muslin. The nipples are conveniently spaced 27 in. apart, center
to center, on the main pipe.

The gas is best introduced at a temperature of 250 deg. F.

248 LEAD SMELTING AND REFINING

Too high a temperature is liable to cause them to ignite. They
are safe at 300 deg. F., but the temperature should not be allowed
to exceed that point.

The gas is cooled by passage through iron pipes of suitable
radiating surface, but the temperature should be controlled by
a dial thermometer close to the bag-house, which should be
observed at least hourly, and there should be an inlet into the
pipe from the outside, so that, in event of rise of temperature
above 300 deg., sufficient cold air may be admitted to reduce it
within the safety limit.

In the case of gas containing much sulphur dioxide, and
especially any appreciable quantity of the trioxide, the bags
should be of unwashed wool. Such gas will soon destroy cotton,
but wool with the natural grease of the sheep still in it is not
much affected. The gas from Scotch hearths and lead-blast
furnaces can be successfully filtered, but the gas from roasting
furnaces contains too much sulphur trioxide to be filtered at all,
bags of any kind being rapidly destroyed.

PART VHI
BLOWERS AND BLOWING ENGINES

ROTARY BLOWERS VS. BLOWING ENGINES FOR
LEAD SMELTING

(April 27, 1901)

A note in the communication from S. E. Bretherton on " Pyritic
Smelting and Hot Blast," published in the Engineering and
Mining Journal of April 13, 1901, refers to a subject of great
interest to lead smelters. Mr. Bretherton remarked that he
had been recently informed by August Raht that by actual
experiment the loss with the ordinary rotary blowers was 100
per cent, under 10 Ib. pressure; that is, it was possible to shut
all the gates so that there was no outlet for the blast to escape
from the blower and the pressure was only 10 Ib., or in other
words the blower would deliver no air against 10 Ib. pressure.
For that reason Mr. Raht expressed himself as being in favor of
blowing engines for lead blast furnaces. This is of special interest,
inasmuch as it comes from one who is recognized as standing in
the first rank of lead-smelting engineers. Mr. Raht is not alone
in holding the opinion he does.

The rotary blower did good service in the old days when the
air was blown into the lead blast furnace at comparatively mod-
erate pressure. At the present time, when the blast pressure
employed is commonly 40 oz. at least, and sometimes as high as
48 oz., the deficiencies of the rotary blower have become more
apparent. Notwithstanding the excellent workmanship which
is put into them by their manufacturers, the extensive surfaces
of contact are inherent to the type, and leakage of air backward
is inevitable and important at the pressures now prevailing.
The impellers of a rotary blower should not touch each other
nor the cylinders in which they revolve, but they are made with
as little clearance as possible, the surfaces being coated with
grease, which fills the clearance space and forms a packing. This
will not, however, entirely prevent leakage, which will naturally
increase with the pressure. Even the manufacturers of rotary
blowers admit the defects of the type, and concede that for pres-

251

252 LEAD SMELTING AND REFINING

sures of 5 Ib. and upward the cylinder blowing engine is the
more economical. Metallurgists are coming generally to the
opinion, however, that blowing engines are probably more eco-
nomical for pressures of 4 Ib. or thereabouts, and some go even
further. With the blowing engines the air-joints of piston and
cylinder are those of actual contact, and the metallurgist may
count on his cubic feet of air, whatever be the pressure. Blowing
engines were actually introduced several years ago by M. W. lies
at what is now the Globe plant of the American Smelting and
Refining Company, and we believe their performance has been
found satisfactory.

The fancied drawback to the use of blowing engines is their
greater first cost, but H. A. Vezin, a mechanical engineer whose
opinions carry great weight, pointed out five years ago in the
Transactions of the American Institute of Mining Engineers
(Vol. XXVI) that per cubic foot of air delivered the blowing
engine was probably no more costly than the rotary blower, but
on the contrary cheaper, stating that the first cost of a cylinder
blower is only 20 to 25 per cent, more than that of a rotary blower
of the same nominal capacity and the engine to drive it. The
capacity of a rotary blower is commonly given as the displace-
ment of the impellers per revolution, without allowance for slip
or leakage backward. Mr. Vezin expressed the opinion that for
the same actual capacity at 2 Ib. pressure, that is, the delivery in
cubic feet against 2 Ib. pressure, the cylinder blower would cost
no more than, if as much as, the rotary blower.

In this connection it is worth while making a note of the
increasing tendency of lead smelters to provide much more pow-
erful blowers than were formerly considered necessary, due, no
doubt, in large measure to the recognition of the greater loss of
air by leakage backward at the pressure now worked against.
It is considered, for example, that a 42 x 140-in. furnace to be
driven under 40-oz. pressure should be provided with a No. 10
blower, which size displaces 300 cu. ft. of air per revolution and
is designed to be run at about 100 r.p.m.; its nominal capacity
is, therefore, 30,000 cu. ft. of air per minute; although its actual
delivery against 40-oz. pressure is much less, as pointed out by
Mr. Raht and Mr. Bretherton. The Connersville Blower Com-
pany, of Connersville, Ind., lately supplied the Aguas Calientes
plant (now of the American Smelting and Refining Company)

BLOWERS AND BLOWING ENGINES 253

with a rotary blower of the above capacity, and duplicates of it
have been installed at other smelting works. The force required
to drive such a huge blower is enormous, being something like
400 h.p., which makes it advisable to provide each blower with
a directly connected compound condensing engine.

In view of the favor with which cylindrical blowing engines
for driving lead blast furnaces are held by many of the leading
lead-smelting engineers, and the likelihood that they will come
more and more into use, it will be interesting to observe whether
the lead smelters will take another step in the tracks of the iron
smelters and adopt the circular form of blast furnace that is
employed for the reduction of iron ore. The limit of size for
rectangular furnaces appears to have been reached in those of
42 x 145 in., or approximately those dimensions. A furnace of
66 x 160 in., which was built several years ago at the Globe
plant at Denver, proved a failure. H. V. Croll at that time
advocated the building of a circular furnace instead of the rect-
angular furnace of those excessive dimensions and considered
that the experience with the latter demonstrated their imprac-
ticability. In the Engineering and Mining Journal of May 28,
1898, he stated that there was no good reason, however, why a
furnace of 300 to 500 tons daily capacity could not be run suc-
cessfully, but considered that the round furnace was the only
form permissible. We are unaware whether Mr. Croll was the
first to advocate the use of large circular furnaces for lead smelt-
ing, but at all events there are other experienced metallurgists
who now agree with him, and the time is, perhaps, not far distant
when they may be adopted.

ROTARY BLOWERS VS. BLOWING ENGINES

BY J. PARKE CHANNING

(June 8, 1901)

In the issues of the Engineering and Mining Journal for
April 13th and 27th reference was made to the relative efficiency
of piston-blowing engines and rotary blowers of the impeller
type, and in these articles August Raht was quoted as saying that,
with an ordinary rotary blower working against 10 Ib. pressure,
the loss was 100 per cent. I have waited some time with the
idea that some of the blower people would call attention to the
concealed fallacy in the statement quoted, but so far have failed
to notice any reference to the matter. I feel quite sure that
Mr. Bretherton failed to quote Mr. Raht in full. The one factor
missing in this statement is the speed at which the blower was
run when the loss was 100 per cent.

The accepted method of testing the volumetric efficiency of
rotary blowers is that of "closed discharge." The discharge
opening of the blower is closed, a pressure gage is connected
with the closed delivery pipe, and the blower is gradually speeded
up until the gage registers the required pressure. The number
of revolutions which the blower makes while holding that pres-
sure, multiplied by the cubic feet per revolution, will give the
total slip of that particular blower at that particular pressure.
Experience has shown that, within the practical limits of speed at
which a blower is run, the slip is a function of the pressure and
has nothing to do with the speed. If, therefore, it were found
that the particular blower referred to by Mr. Raht were obliged
to be revolved at the rate of 30 r.p.m. in order to maintain a
constant pressure of 10 Ib. with a closed discharge, and if the
blower were afterward put in practical service, delivering air,
and were run at a speed of 150 r.p.m., it would then follow that
its delivery of air would amount to: 150 — 30 = 120. Its volu-
metric efficiency would be 120 -r- 150 = 80 per cent. The above

254

BLOWERS AND BLOWING ENGINES 255

figures must not be relied upon, as I give them simply by way
of illustration.

About a year ago I had the pleasure of examining the tabu-
lated results of some extensive experiments in this direction,
made by one of the blower companies. I believe they carried
their experiments up to 10 Ib. pressure, and I regret that I have
not the figures before me, so that I could give something definite.
I do, however, remember that in the experimental blower, when
running at about 150 r.p.m., the volumetric efficiency at 2 Ib.
pressure was about 85 per cent., and that at 3 Ib. pressure the
volumetric efficiency was about 81 per cent.

It is unnecessary in this connection to call attention to the
horse-power efficiency of rotary blowers. This is a matter entirely
by itself, and there is considerable difference of opinion among
engineers as to the relative horse-power efficiency of rotary
blowers and piston blowers. All agree that there is a certain
pressure at which the efficiency of the blower becomes less than
the efficiency of the blowing engine. This I have heard placed
all the way from 2 Ib. up to 6 Ib.

At the smelting plant of the Tennessee Copper Company we
have lately installed blast-furnace piston-blowing engines; the
steam cylinders are of the Corliss type and are 13 and 24 in. by
42 in.; the blowing cylinders are two in number, each 57 x42 in.;
the air valves are all Corliss in type. These blowing engines are
designed to operate at a maximum air pressure of 2J Ib. per
square inch.

At the Santa Fe Gold and Copper Mining Company's smelter
we have recently installed a No. 8 blower directly coupled to a
14 x 32-in. Corliss engine. This blower has been in use about
five months and is giving very good results against the compara-
tively low pressure of 12 oz., or } Ib.

During the coming summer it is my intention to make careful
volumetric and horse-power tests on these two types of machines
under similar conditions of air pressure, and to publish the
results; but in the meantime I wish to correct the error that a
rotary blower of the impeller type is not a practicable machine
at pressure over 5 Ib.

BLOWERS AND BLOWING ENGINES FOR LEAD AND
COPPER SMELTING

BY HIRAM W. HIXON

(July 20, 1901)

In the Engineering and Mining Journal for July 6th I note
the discussion over the relative merits of blowers and blowing
engines for lead and copper smelting, and wish to state that,
irrespective of the work to be done, the blast pressure will depend
entirely on the charge burden in any kind of blast-furnace work,
and that the charge burden governs the reducing action of the
furnace altogether. Along these lines the iron industry has
raised the charge burden up to 100 ft. to secure the full benefit
of the reducing action of the carbon monoxide on the ore.

In direct opposition to this we have what is known as pyritic
smelting, wherein the charge burden governs the grade of the
matte produced to such an extent that if a charge run with
4 to 6 ft. of burden above the tuyeres, producing 40 per cent,
matte, is changed to a charge burden of 10 or 12 ft., the grade
of the matte will decrease from 40 per cent, to probably less than
20 per cent. I can state this as a fact from recent experience in
operating a blast furnace on heap-roasted ores under the condi-
tions named, with the result as above stated.

I was exceedingly skeptical about pyritic smelting as advo-
cated by some of your correspondents, and still continue to be
so; but on making inquiries from some of my co-workers in this
line, Mr. Sticht of Tasmania, and Mr. Nutting of Bingham,
Utah, I have arrived at the following conclusion, to which some
may take exception: That pyritic smelting without fuel, or with
less than 5 per cent., with hot blast, is practically impossible;
that smelting raw ore with a low charge burden to avoid the
reducing action of the carbon monoxide, thereby securing oxida-
tion of the iron and sulphur, is possible and practicable, under
favorable conditions; and that a large portion of the sulphur is
burned off, and the iron, without reducing action, goes into the

256

BLOWERS AND BLOWING ENGINES 257

slag in combination with silica. These results can be obtained
with cold blast.

A blowing engine would certainly be much out of place for
operating copper-matting furnaces run with the evident intention
of oxidizing sulphur and iron and securing as high a grade of
matte as possible, for the reason that to do this it is necessary
to run a low charge burden, and with a low charge burden a
high pressure of blast cannot be maintained. With a 4- to 6-ft.
charge burden the blast pressure at Victoria Mines at present is
3 oz., produced by a No. 6 Green blower run at 120 r.p.m.; and
a blowing engine, delivering the same amount of air, would cer-
tainly not give more pressure. Under the conditions which we
have, a fan would be more effective than a pressure blower, and
a blowing engine entirely out of the question as far as economy is
concerned.

I installed blowing engines at the East Helena for lead smelting
where the charge burden was 21 ft. and the blast pressure at
times went up as high as 48 oz. Under these conditions the
blowing engines gave satisfaction, but I am of the opinion that
the same amount of blast could have been obtained under that
pressure with less horse-power by the best type of rotary blower.
I do not believe that the field of the blowing engine properly
exists below 5 lb., and if this pressure cannot be obtained by
charge-burden conditions, their installation is a mistake.

I wish to add the very evident fact that varying the grade
of the matte by feeding the furnace at different hights varies
the slag composition as to its silica and iron contents and makes
the feeder the real metallurgist. The reducing action in the
furnace is effected almost entirely by the gases, and when these
are allowed to go to waste, reduction ceases.

BLOWING ENGINES AND ROTARY BLOWERS — HOT
BLAST FOR PYRITIC SMELTING

BY S. E. BRETHERTON

(August 24, 1901)

I have just read in the Engineering and Mining Journal of
July 20th an interesting letter written by Hiram W. Hixon in
regard to blowing engines versus the rotary blowers, and also
the use of cold blast for pyritic smelting.

The controversy, which I unintentionally started in my letter
in the Engineering and Mining Journal of April 13th last, about
the advantages of using either blowers or blowing engines for
blast furnaces, does not particularly interest me, with the excep-
tion that I have about decided, in my own mind, to use blowing
engines where there is much back pressure, and the ordinary
up-to-date blower for pyritic or matte smelting where much back
pressure should not exist. I fully appreciate the fact that so-
called pyritic smelting can be done to a limited extent, even
with cold blast. Theoretically, enough oxygen can be sent into
the blast furnace, contained in the cold blast, to oxidize both
the fuel and the sulphur in an ordinary sulphide charge, but I
have not yet learned where a high concentration is being made
with unroasted ore and cold blast. I experimented on these
lines at different times for three years, during 1896, 1897, and
1898, making a fair concentration with refractory ores, most of
which had been roasted. I was myself interested in the profits
and as anxious as any one for economy. We tried, for fuel in
the blast furnace, coke alone, coke and lignite coal, lignite coal
alone, lignite coal and dry wood, coal and green wood, and then
coke and green wood, all under different hights of ore burden
in the furnace.

A description of these experiments would, no doubt, be tire-
some to your readers, but I wish to state that the furnace was
frozen up several times on account of using too little fuel, when
the cold blast would gradually drive nearly all the heat to the

258

BLOWERS AND BLOWING ENGINES 259

top of the furnace, the crucible and between the tuyeres becoming
so badly crusted that the furnace had to be cleaned out and
blown in again, unless I was called in time to save it by changing
the charge and increasing the fuel. We were making high-grade
matte under contract, high concentration and small matte fall,
which would, no doubt, aggravate matters.

After the introduction of hot blast, heated up to between
200 and 300 deg. F., we made the same grade of matte from the
same character of ore, with the exception that we then smelted
without roasting, and reduced the percentage of fuel consump-
tion, increased the capacity of the furnace, and almost entirely
obviated the trouble of cold crucibles and hot tops. I write the
above facts, as they speak for themselves.

I nearly agree with Mr. Hixon, and do not think it practical
to smelt with much less than 5 per cent, coke continuously; but
there is a great saving between the amount of coke used with a
moderately heated blast and cold blast. Regardless of either
hot or cold blast, however, the fuel consumption depends very
much on the character of the ore to be smelted, the amount of
matte-fall and grade of matte made. It is not always advisable
or necessary to use hot blast for a matting furnace; that is, where
the supply of sulphur is limited. It may then be necessary to
use as much fuel in the blast furnace to prevent the sulphur
from oxidizing as will be sufficient to furnish the heat for smelting.
Such conditions existed at Silver City, N. M., at times, after our
surplus supply of iron and zinc sulphide concentrates was used.
I understand that they are now short of sulphur there, on account
of getting a surplus amount of oxidized copper ore, and are only
utilizing what little heat the slag gives them, without the addition
of any fuel on top of the forehearth.

Before closing this, which I intended to have been brief, I
wish to call your attention to a little experience we had with
-alumina in the matting furnace at Silverton, Col., where I was
acting as consulting metallurgist. The ore we had to smelt
contained, on an average, about 20 per cent. A12O3, 30 per cent.
Si02, with 18 per cent. Fe in the form of an iron pyrite, and no
other iron was available except some iron sulphide concentrates
containing a small percentage of zinc and lead.

The question naturally arose, could we oxidize and force
sufficient of the iron into the slag, and where should we class

260 LEAD SMELTING AND REFINING

the alumina, as a base or an acid? My experience in lead smelting
led me to believe that A12O3 could only be classed as an acid in
the ordinary lead furnace, and that it would be useless to class
it otherwise in a shallow matting furnace; and E. W. Walter,
the superintendent and metallurgist in charge, agreed with me.

We then decided to make a bisilicate slag, classing the alumina
as silica, and we obtained fairly satisfactory results. The slag
made was very clean, but treacherous, which was attributed to
two reasons: First, that it required more heat to keep the alumina
slag liquid enough to flow than it does a nearly straight silica
slag; and, second, that we were running so close to the formula
of a bisilicate and aluminate slag (about 31 J per cent. SiO2,
27 per cent. Fe, 20 per cent. CaO, and 18 per cent. A12O3, or
49 J per cent, acid) that a few charges thrown into the furnace
containing more silica or alumina than usual would thicken the
slag so that it would then require some extra coke and flux to
save the furnace. At times the combined SiO2 and A12O3 did
reach 55 and 56 per cent, in the slag, which did not freeze up
the furnace, but caused us trouble.

PART IX
LEAD REFINING

THE REFINING OF LEAD BULLION1

BY F. L. PlDDINGTON
(October 3, 1903)

In presenting this account of the Parkes process of desilver-
izing and refining lead bullion no originality is claimed, but I
hope that a description of the process as carried out at the works
of the Smelting Company of Australia may be of service.

Introductory. — The Parkes process may be conveniently
summarized as follows:

1. Softening of the base bullion to remove copper, antimony,
etc.

2. Removal of precious metals from the softened bullion by
means of zinc.

3. Refining the desilverized lead.

4. Liquation of gold and silver crusts obtained from operation
No. 2.

5. Retorting the liquated alloy to drive off zinc.

6. Concentrating and refining bullion from No. 5.
Softening. — This is done in reverberatory furnaces. In large

works two furnaces are used, copper, antimony, and arsenic being
removed in the first and antimony in the second. The size of
the furnaces is naturally governed by the quantity to be treated.
In these works (refining about 200 tons weekly) a double set of
15-ton furnaces were at work. The sides and ends of these
furnaces are protected by a jacket with a 2-in. water space, the
jacket extending some 3 in. above the charge level and 6 in. to
9 in. below it. The furnace is built into a wrought-iron pan,
and if the brickwork is well laid into the pan there need be no
fear of lead breaking through below the jacket.

The bars of bullion (containing, as a rule, 2 to 3 per cent, of
impurities) are placed in the furnace carefully, to avoid injuring
the hearth, and melted down slowly. The copper dross separates

1 Abstract from the Journal of the Chemical, Metallurgical and Mining
Society of South Africa, May, 1903.

263

264 LEAD SMELTING AND REFINING

out and floats on top of the charge, which is stirred frequently
to expose fresh surfaces. If the furnace is overheated some
dross is melted into the lead again and will not separate out
until the charge is cooled back. However carefully the work is
done some copper remains with the lead, and its effects are to
be seen in the later stages. The dross is skimmed into a slag
pot with a hole bored in it some 4 in. from the bottom; any lead
drained from the pot is returned to the charge. The copper
dross is either sent back to the blast furnace direct or may be
first liquated. By the latter method some 30 per cent, of the
lead contents of the dross is recovered in the refinery.

Base bullion made at a customer's smelter will often vary
greatly in composition, and it is, therefore, difficult to give any
hard and fast figures as to percentage of metals in the dross.
As a rule our dross showed 65 to 70 per cent, lead, copper 2 to
9 per cent, (average 4 per cent.), gold and silver values varying
with the grade of the original bullion, though it was difficult to
detect any definite relation between bullion and dross. It was,
however, noticed that gold and silver values increased with the
percentage of copper.

Immediately the copper dross is skimmed off the heat is
raised considerably, and very soon a tin (and arsenic, if present)
skimming appears. It is quite "dry" and may be removed in
an hour or so. It is a very small skimming, and the tin, not
being worth saving, is put with the copper dross.

The temperature is now raised again and antimony soon
shows in black, boiling, oily drops, gathering in time into a sheet
covering the surface of the lead. When the skimming is about
i-inch thick, slaked lime, ashes, or fine coal is thrown on and
stirred in. The dross soon thickens up and may be skimmed
off easily. This operation is repeated until all antimony is elim-
inated. Constant stirring of the charge is necessary. The
addition of litharge greatly facilitates the removal of antimony;
either steam or air may be blown on the surface of the metal to
hasten oxidation, though they have anything but a beneficial
effect on the furnace lining. From time to time samples of the
dross are taken in a small ladle, and after setting hard the sample
is broken in two. A black vitreous appearance indicates plenty of
antimony yet in the charge. Later samples will look less black,
until finally a few yellowish streaks are seen, being the first

LEAD REFINING 265

appearance of litharge. When all antimony is out the fracture
of a sample should be quite yellow and the grain of the litharge
long, a short grain indicating impurities still present, in which
case another skimming is necessary. The analysis of a represen-
tative sample of antimony dross was as follows:

PbO = 78. 1 1 per cent. CaO =1.10 per cent.

Sb2O4 = 8.75 " " Fe2O3 = 0.42 " "

As2O3 = 2.18 " " AJ2O3 = 0.87 " "

CuO= 0.36 " " Insol. = 4.10" "

Antimony dross is usually kept separate and worked up from
time to time, yielding hard antimonial lead, used for type metal,
Britannia metal, etc.

Desilverization. — The softening being completed the charge
is tapped and run to a kettle or pan of cast iron or steel, holding,
when conveniently full, some 12 or 13 tons. The lead falling
into the kettle forms a considerable amount of dross, which is
skimmed off and returned to the softening furnace. By cooling
down the charge until it nearly "freezes" an additional copper
skimming is obtained, which also is returned to the softener. The
kettle is now heated up to the melting point of zinc, and the
zinc charge, determined by the gold and silver contents of the
kettle, is added and melted. The charge is stirred, either by
hand or steam, for about an hour, after which the kettle is allowed
to cool down for some three hours and the first zinc crust taken
off. When the charge is skimmed clean a sample of the bullion
is taken for assay, and while this is being done the kettle is heated
again for the second zinc charge, which is worked in the same
way as the first; sometimes a third addition of zinc is necessary.
The resulting crusts are kept separate, the second and third being
added to the next charge as "returns," allowing 3 Ib. of zinc in
returns as equal to 1 Ib. of fresh zinc. An alternative method
is to take out gold and silver in separate crusts, in which case
the quantity of zinc first added is calculated on the gold contents
of the kettle only. The method of working is the same, though
subsequent treatment may differ in that the gold crusts are
cupeled direct.

As to the quantity of zinc required:

1. Extracting the gold with as little silver as possible, the
following figures were obtained:

266

LEAD SMELTING AND REFINING

Total Gold— Au.

In kettle 300 oz. 1 Ib. zinc takes out 1.00 oz.

" " 200 " " " " " 1.00 "

" " 150 " " " " " 0.79 "

" " 100 " " " " " 0.59 "

" 60 " " " " " 0.45 "

2. Silver zinking gave the following general results with
11-ton charges:

Total Silver—

In kettle ................. 1,450 oz.

1,200 "

930 "

J"!'!!^;*!"] 755 "

616 "

460 "

" "
" "

1 Ib. zinc takes out 5.6 oz.

II O Q II

II 35 II

II n A l<

i« r> c «

ZJO

3. Extracting gold and silver together:

Au. Oz.

Ac. Oz.

Au. Oz.

Ao. Oz.

494

3,110

0.59

3.60

443

1,883

0.64

2.80

330

2,417

0.45

3.34

204

1,638

0.36

2.86

143

1,330

0.28

2.65

123

1,320

0.23

2.54

It will be noticed that in each case the richer the bullion the
greater the extractive power of zinc. Experiments made on
charges of rich bullion showed that the large amount of zinc
called for by the table in use was unnecessary, and 250 Ib. was
fixed on as the first addition of zinc. On this basis an average
of 237 charges gave results as follows:

• TOTAL C
Au. Oz.

ZINC USED
LBS.

IT w 7rjjr

TAKES Our .
Ao. Oz.

Ac. Oz.

Au. Oz.

520

1,186

507.5

1.27

2.91

The zinc used was that necessary to clean the kettle, added
as follows: 1st, 250 Ib.; 2d (average), 127 Ib.; 3d (average), 57 Ib.
In 112 cases no third addition was required. From these figures
it appears that in the earlier work the zinc was by no means
saturated.

LEAD REFINING 267

Refining the Lead. — Gold and silver being removed, the lead
is siphoned off into the refining kettle and the fire made up.
In about four hours the lead will be red hot, and when hot enough
to burn zinc, dry steam, delivered by a J-in. pipe reaching nearly
to the bottom of the kettle, is turned on. The charge is stirred
from time to time and wood is fed on the top to assist dezinking
and prevent the formation of too much litharge. In three to
four hours the lead will be soft and practically free from zinc.
When test strips show the lead to be quite soft and clean, the
kettle is cooled down and the scum of lead and zinc oxides skimmed
off. In an hour or so the lead will be cool enough for molding;
the bar should have a yellow luster on the face when set; if the
lead is too cold it will be white, if too hot a deep blue. The
refining kettles are subjected to severe strain during the steaming
process, and hence their life is uncertain — an average would be
about 60 charges; the zinking kettles, on the other hand, last
very much longer. Good steel kettles (if they can be obtained)
are preferable to cast iron.

Treatment of Zinc Crusts. — Having disposed of the lead, let
us return now to the zinc crusts. These are first liquated in a
small reverberatory furnace, the hearth of which is formed of a
cast-iron plate (the edges of the long sides being turned up some
4 in.) laid on brasque filling, with a fall from bridge to flue of
| in. per foot and also sloping from sides to center. The opera-
tion is conducted at a low temperature and the charge is turned
over at intervals, the liquated lead running out into a small
separately fired kettle. This lead rarely contains more than a
few ounces of silver per ton; it is baled into bars, and returned to
the zinking kettles or worked up in a separate charge. In two
to three hours the crust is as "dry" as it is advisable to make
it, and the liquated alloy is raked out over a slanting perforated
plate to break it up and goes to the retort bin.

Retorting the Alloy. — This is carried on in Faber du Faur
tilting furnaces — simply a cast-iron box swinging on trunnions
and lined with firebrick. Battersea retorts (class 409) holding
560 Ib. each are used; their average life is about 30 charges.
The retorts are charged hot, a small shovel of coal being added
with the alloy. The condenser is now put in place and luted on;
it is made of J-in. iron bent to form a cylinder 12 in. in diameter,
open at one end; it is lined with a mixture of lime, clay and

268 LEAD SMELTING AND REFINING

cement. It has three holes, one on the upper side close to the
furnace and through which a rod can be passed into the retort,
a vent-hole on the upper side away from the furnace, and a tap-
hole on the bottom for condensed zinc. In an hour or so the
flame from the vent-hole should be green, showing that distillation
has begun. When condensation ceases (shown by the flame) the
condenser is removed and the bullion skimmed and poured into
bars for the cupel. The products of retorting are bullion, zinc,
zinc powder and dross. Bullion goes to the cupel, zinc is used
again in the desilverizing kettles, powder is sieved to take out
scraps of zinc and returned to the blast furnace, or it may be,
and sometimes is, used as a precipitating agent in cyanide works;
dross is either sweated down in a cupel with lead and litharge,
together with outside material such as zinc gold slimes from
cyanide works, jeweler's sweep, mint sweep, etc., or in the soften-
ing furnace after the antimony has been taken off. In either
case the resulting slag goes back to the blast furnace. The total
weight of alloy treated is approximately 7 per cent, of the original
base bullion. The zinc recovered is about 60 per cent, of that
used in desilverizing. The most important source of temporary
loss is the retort dross (consisting of lead-zinc-copper alloy with
carbon, silica and other impurities), and it is here that the neces-
sity of removing copper in the softening process is seen, since any
copper comes out with the zinc crusts and goes on to the retorts,
where it enters the dross, carrying gold and silver with.it. If
much copper is present the dross may contain more gold and silver
than the retort bullion itself. In this connection I remember an
occasion on which some retort dross yielded gold and silver to
the extent of over 800 and 3000 oz. per ton respectively.

Cupellation. — Retort bullion is first concentrated (together
with bullion resulting from dross treatment) to 50 to 60 per cent,
gold and silver in a water-jacketed cupel. The side lining is
protected by an inch water-pipe imbedded in the lining at the
litharge level or by a water-jacket, the inner face of which is of
copper; the cupel has also a water- jacketed breast so that the
front is not cut down. The cupel lining may be composed of
limestone, cement, fire-clay and magnesite in various proportions,
but a simple lining of sand and cement was found quite satis-
factory. When the bullion is concentrated up to 50 to 60 per
Cent, gold and silver, it is baled out and transferred to the finishing

LEAD REFINING 269

cupel, where it is run up to about 0.995 fine; it is then ready either
for the melting-pot or parting plant. The refining test, by the
way, is not water-cooled.

Re-melting is done in 200-oz. plumbago crucibles and presents
no special features. In the case of dore bullion low in gold,
"sprouting" of the silver is guarded against by placing a piece
of wood or charcoal on the surface of the metal before pouring,
and any slag is kept back. The quantity of slag formed is, of
course, very small, so that the bars do not require much cleaning.

The parting plant was not in operation in my time, and I
am therefore unable to go into details. The process arranged
for was briefly as follows: Solution of the dore bullion in H2SO4;
crystallization of silver as monosulphate by dilution and cooling;
decomposition of silver sulphate by ferrous sulphate solution
giving metallic silver and ferric sulphate, which is reduced to
the ferrous salt by contact with scrap iron. The gold and silver
are washed thoroughly with hot water and cast into bars.

In conclusion, some variations in practice may be noted.
The use of two furnaces in the softening process has already
been mentioned; by this means the dressing and softening are
more perfect and subsequent operations thereby facilitated;
further, the furnaces, being kept at a more equable temperature,
are less subject to wear and tear. Zinc crusts are sometimes
skimmed direct into an alloy press in which the excess of lead is
squeezed out while still molten; liquation is then unnecessary.
Refining of the lead may be effected by a simple scorification in
a reverberatory, the soft lead being run into a kettle from which
it is molded into market bars.

THE ELECTROLYTIC REFINING OF BASE LEAD
BULLION

BY TITUS ULKE

(October 11, 1902)

Important changes in lead-refining practice are bound to
follow, in my opinion, the late demonstration on a large scale
of the low working cost and high efficiency of Betts' electrolytic
process of refining lead bullion. It was my good fortune recently
to see this highly interesting process in operation at Trail, British
Columbia, through the kindness of the inventor, A. G. Betts,
and Messrs. Labarthe and Aldridge, of the Trail works.

A plant of about 10 tons daily capacity, which probably cost
about $25,000, although it could be duplicated for perhaps
$15,000 at the present time, was installed near the Trail smelting
works. It has been in operation for about ten months, I am
informed, with signal success, and the erection of a larger plant,
of approximately 30 tons capacity and provided with improved
handling facilities, is now completed.

The deposit ing-room contains 20 tanks, built of wood, lined
with tar, and approximately of the size of copper-refining tanks.
Underneath the tank-room floor is a basement permitting inspec-
tion of the tank bottoms for possible leakage and removal of the
solution and slime. A suction pump is employed in lifting the
electrolyte from the receiving tank and circulating the solution.
In nearly every respect the arrangement of the plant and its
equipment is strikingly like that of a modern copper refinery.

The great success of the process is primarily based upon
Betts' discovery of the easy solubility of lead in an acid solution
of lead fluosilicate, which possesses both stability under elec-
trolysis and high conductivity, and from which exceptionally pure
lead may be deposited with impure anodes at a very low cost.
With such a solution, there is no polarization from formation of
lead peroxide on the anode, no evaporation of constituents
except water, and no danger in its handling. It is cheaply

270

LEAD REFINING 271

obtained by diluting hydrofluoric acid of 35 per cent. HF, which
is quoted in New York at 3c. per pound, with an equal volume
of water and saturating it with pulverized quartz according to
the equation:

Si02 + 6HF = H SiF6 + 2H20.

According to Mr. Betts, an acid of 20 to 22 per cent, will
come to about $1 per cu. ft., or to $1.25 when the solution has
been standardized with 6 Ib. of lead. One per cent, of lead will
neutralize 0.7 per cent. H2SiFe. The electrolyte employed at
the time of my inspection of the works contained, I believe,
8 per cent, lead and 11 per cent, excess of fluosilicic acid.

The anodes consist of the lead bullion to be refined, cast into
plates about 2 in. thick and approximately of the same size as
ordinary two-lugged copper anodes. Before being placed in
position in the tanks, they are straightened by hammering over
a mold and their lugs squared. No anode sacks are employed
as in the old Keith process.

The cathode sheets which receive the regular lead deposits
are thin lead plates obtained by electrodeposition upon and
stripping from special cathodes of sheet steel. The latter are
prepared for use by cleaning, flashing with copper, lightly lead-
plating in the tanks, and greasing with a benzine solution of
paraffin, dried on, from which the deposited lead is easily stripped.

The anodes and cathodes are separated by a space of 1J to
2 in. in the tank and are electrically connected in multiple, the
tanks being in series circuit. The fall in potential between
tanks is only about 0.2 of a volt, which remarkably low voltage
is due to the high conducting power of the electrolyte and to some
extent to the system of contacts used. These contacts are small
wells of mercury in the bus-bars, large enough to accommodate
copper pins soldered to the iron cathodes or clamped to the
anodes. Only a small amount of mercury is required.

Current strengths of from 10 to 25 amperes per sq. ft.
have been used, but at Trail 14 amperes have given the most
satisfactory results as regards economy of working and the
physical and chemical properties of the refined metal produced.

A current of 1 ampere deposits 3.88 grams of lead per hour,
or transports 3J times as much lead, in this case, as copper with
an ordinary copper-refining solution. A little over 1000 kg.,: or

272 LEAD SMELTING AND REFINING

2240 lb., requires about 260,000 ampere hours. At 10 amperes
per sq. ft. the cathode (or anode) area should be about 1080
sq. ft. per ton of daily output. Taking a layer of electrolyte
1.5 in. thick, 135 cu. ft. will be found to be the amount between
the electrodes, and 175 cu. ft. may be taken as the total quantity
of solution necessary, according to Mr. Betts' estimate. The
inventor states that he has worked continuously and successfully
with a drop of potential of only 0.175 volt per tank, and that
therefore 0.25 volt should be an ample allowance in regular
refining. Quoting Mr. Betts: " 260,000 ampere hours at 0.25
volt works out to 87 electrical h.p. hours of 100 h.p. hours at
the engine shaft, in round numbers. Estimating that 1 h.p.
hour requires the burning of 1.5 lb. of coal, and allowing say
60 lb. for casting the anodes and refined lead, each ton of lead
refined requires the burning of 210 lb. of fuel." With coal at
$6 per ton the total amount of fuel consumed, therefore, should
not cost over 60c., which is far below the cost of fire-refining
base lead bullion, as we know.

In the Betts electrolytic process, practically all the impurities
in the base bullion remain as a more or less adherent coating
on the anode, and only the zinc, iron, cobalt and nickel present
go into solution. The anode residue consists practically of all
the copper, antimony, bismuth, arsenic, silver and gold con-
tained in the bullion, and very nearly 10 per cent, of its weight
in lead. Having the analysis of any bullion, it is easy to calcu-
late with these data the composition of the anode residue and
the rate of pollution of the electrolyte. Allowing 175 cu. ft. of
electrolyte per ton of daily output, it will be found that in the
course of a year these impurities will have accumulated to the
extent of a very few per cent. Estimating that the electrolyte
will have to be purified once a year, the amount to be purified
daily is less than 1 cu. ft. for each ton of output. The amount
of lead not immediately recovered in pure form is about 0.3 per
cent., most of which is finally recovered. As compared with the
ordinary fire-refined lead, the electrolytically refined lead is much
purer and contains only mere traces of bismuth, when bismuthy
base bullion is treated. Furthermore, the present loss of silver
in fire refining, amounting, it is claimed, to about 1J per cent, of
the silver present, and covered by the ordinary loss in assay, is to
a large extent avoided, as the silver in the electrolytic process is

LEAD REFINING 273

concentrated in the anode residue with a very small loss, and
the loss of silver in refining the slimes is much less than in treating
the zinc crusts and refining the silver residue after distillation.
The silver slimes obtained at Trail, averaging about 8000 oz. of
gold and silver per ton, are now treated at the Seattle Smelting
and Refining Works. There the slimes are boiled with concen-
trated sulphuric acid and steam, allowing free access of air,
which removes the greater part of the copper. The washed
residue is then dried in pans over steam coils, and melted down
in a magnesia brick-lined reverberatory, provided with blast
tuyeres, and refined. In this reverberatory furnace the remainder
of the copper left in the slimes after boiling is removed by the
addition of niter as a flux, and the antimony with soda. The
dore bars finally obtained are parted in the usual way with
sulphuric acid, making silver 0.999 fine and gold bars at least
0.992 fine.

Mr. Betts treated 2000 grams of bullion, analyzing 98.76 per
cent. Pb, 0.50 Ag, 0.31 Cu, and 0.43 Sb with a current of 25
amp. per square foot in an experimental way, and obtained
products of the following composition:

Refined Lead: 99.9971 per cent. Pb, 0.0003 Ag, 0.0007 Cu,
and 0.0019 Sb.

Anode Residue: 9.0 per cent. Pb, 36.4 Ag, 25.1 Cu, and 2.95 Sb.

Four hundred and fifty pounds of bullion from the Compania
Metalurgica Mexicana, analyzing 0.75 per cent. Cu, 1.22 Bi,
0.94 As, 0.68 Sb, and assaying 358.9 oz. Ag and 1.71 oz. Au per
ton, were refined with a current of 10 amp. per square foot, and
gave a refined lead of the following analysis: 0.00027 per cent.
Cu, 0.0037 Bi, 0.0025 As, 0 Sb, 0.0010 Ag, 0.0022 Fe, 0.0018
Zn and Pb (by difference) 99.9861 per cent.

Although the present method for recovering the precious
metals and by-products from the anode residue leaves much
room for improvement, the use of the Betts process may be
recommended to our lead refiners, because it is a more economical
and efficient method than the fire-refining process now in common
use. I will state my belief, in conclusion, that the present devel-
opment of electrolytic lead refining signalizes as great an advance
over zinc desilverization and the fire methods of refining lead as
electrolytic copper refining does over the old Welsh method of
refining that metal.

ELECTROLYTIC LEAD-REFINING J

BY ANSON G. BETTS

A solution of lead fluosilicate, containing an excess of fluo-
silicic acid, has been found to work very satisfactorily as an
electrolyte for refining lead. It conducts the current well, is
easily handled and stored, non-volatile and stable under elec-
trolysis, may be made to contain a considerable amount of dis-
solved lead, and is easily prepared from inexpensive materials.
It possesses, however, in common with other lead electrolytes,
the defect of yielding a deposit of lead lacking in solidity, which
grows in crystalline branches toward the anodes, causing short
circuits. But if a reducing action (practically accomplished by
the addition of gelatine or glue) be given to the solution, a per-
fectly solid and dense deposit is obtained, having very nearly
the same structure as electrolytically deposited copper, and a
specific gravity of about 11.36, which is that of cast lead.

Lead fluosilicate may be crystallized in very soluble bril-
liant crystals, resembling those of lead nitrate and containing
four molecules of water of crystallization, with the formula
PbSiF6,4H20. This salt dissolves at 15 deg. C. in 28 per cent,
of its weight of water, making a syrupy solution of 2.38 sp. gr.
Heated to 60 deg. C., it melts in its water of crystallization. A
neutral solution of lead fluosilicate is partially decomposed on
heating, with the formation of a basic insoluble salt and free
fluosilicic acid, which keeps the rest of the salt in solution. This
decomposition ends when the solution contains perhaps 2 per
cent, of free acid; and the solution may then be evaporated
without further decomposition. The solutions desired for re-
fining are not liable to this decomposition, since they contain
much more than 2 per cent, of free acid. The electrical con-
ductivity depends mainly on the acidity of the solution.

My first experiments were carried out without the addition

1 Abstract of a paper in Transactions American Institute of Mining
Engineers, XXXIV (1904), p. 175.

274

LEAD REFINING 275

of gelatine to the fluosilicate solution. The lead deposit con-
sisted of more or less separate crystals that grew toward the
anode, and, finally, caused short circuits. The cathodes, which
were sheet-iron plates, lead-plated and paraffined, had to be
removed periodically from the tanks and passed through rolls,
to pack down the lead. When gelatine has been added in small
quantities, the density of the lead is greater than can be produced
by rolling the crystalline deposit, unless great pressure is used.

The Canadian Smelting Works, Trail, B. C., have installed a
refinery, making use of this process. There are 28 refining-tanks,
each 86 in. long, 30 in. wide and 42 in. deep, and each receiving
22 anodes of lead bullion with an area of 26 by 33 in. exposed to
the electrolyte on each side, and 23 cathodes of sheet lead, about
JL in. thick, prepared by deposition on lead-plated and paraffined
iron cathodes. The cathodes are suspended from 0.5 by 1 in.
copper bars, resting crosswise on the sides of the tanks. The
experiment has been thoroughly tried of using iron sheets to
receive a deposit thicker than -^ in.; that is, suitable for direct
melting without the necessity of increasing its weight by further
deposition as an independent cathode; but the iron sheets are
expensive, and are slowly pitted by the action of the acid solu-
tion; and the lead deposits thus obtained are much less smooth
and pure than those on lead sheets.

The smoothness and the purity of the deposited lead are
proportional. Most of the impurity seems to be introduced
mechanically through the attachment of floating particles of
slime to irregularities on the cathodes. The effect of roughness
is cumulative: it is often observed that particles of slime attract
an undue amount of current, resulting in the lumps seen in the
cathodes. Samples taken at the same time showed from 1 to
2.5 oz. silver per ton in rough pieces from the iron cathodes, 0.25
oz. as an average for the lead-sheet cathodes, and only 0.04 oz.
in samples selected for their smoothness. The variation in the
amount of silver (which is determined frequently) in the samples
of refined lead is attributed not to the greater or less turbidity
of the electrolyte at different times, but to the employment of
new men in the refinery, who require some experience before
they remove cathodes without detaching some slime from the
neighboring anodes.

Each tank is capable of yielding, with a current of 4000

276 LEAD SMELTING AND REFINING

amperes, 750 Ib. of refined lead per day. The voltage required to
pass this current was higher than expected, as explained below;
and for this reason, and also because the losses of solution were
very heavy until proper apparatus was put in to wash thoroughly
the large volume of slime produced (resulting in a weakened
electrolyte), the current used has probably averaged about 3000
amperes. The short circuits were also troublesome, though this
difficulty has been greatly reduced by frequent inspection and
careful placing of the electrodes. At one time, the solution in
use had the following composition in grams per 100 c.c.: Pb,
6.07; Sb, 0.0192; Fe, 0.2490; SiF6, 6.93, and As, a trace. The
current passing was 2800 amperes, with an average of about
0.44 volts per tank, including bus-bars and contacts. It is not
known what was the loss of efficiency on that date, due to short
circuits; and it is, therefore, impossible to say what resistance
this electrolyte constituted.

Hydrofluoric acid of 35 per cent., used as a starting material
for the preparation of the electrolyte, is run by gravity through
a series of tanks for conversion into lead fluosilicate. In the
top tank is a layer of quartz 2 ft. thick, in passing through which
the hydrofluoric acid dissolves silica, forming fluosilicic acid.
White lead (lead carbonate) in the required quantity is added in
the next tank, where it dissolves readily and completely with
effervescence. All sulphuric acid and any hydrofluoric acid that
may not have reacted with silica settle out in combination with
lead as lead sulphate and lead fluoride. Lead fluosilicate is one
of the most soluble of salts; so there is never any danger of its
crystallizing out at any degree of concentration possible under
this method. The lead solution is then filtered and run by gravity
into the refining-tanks.

The solution originally used at Trail contained about 6 per
cent. Pb and 15 per cent. SiF6.

The electrical resistance in the tanks was found to be greater
than had been calculated for the same solution, plus an allow-
ance for loss of voltage in the contacts and conductors. This
is partly, at least, due to the resistance to free motion of the
electrolyte, in the neighborhood of the anode, offered by a layer
of slime which may be anything up to i in. thick. During elec-
trolysis, the SiF6 ions travel toward the anodes, and there com-
bine with lead. The lead and hydrogen travel in the opposite

LEAD REFINING 277

•direction and out of the slime; but there are comparatively few
lead ions present, so that the solution in the neighborhood of
the anodes must increase in concentration and tend to become
neutral. This greater concentration causes an e.m.f. of polar-
ization to act against the e.m.f. of the dynamo. This amounted
to about 0.02 volt for each tank. The greater effect comes from
the greater resistance of the neutral solution with which the
slime is saturated. There is, consequently, an advantage in
working with rather thin anodes, when the bullion is impure
enough to leave slime sticking to the plates. A compensating
advantage is found in the increased ease of removing the slime
with the anodes, and wiping it off the scrap in special tanks,
instead of emptying the tanks and cleaning out, as is done in
copper refineries.

It is very necessary to have adequate apparatus for washing
solution out of the slime. The filter first used consisted of a
supported filtering cloth with suction underneath. It was very
difficult to get this to do satisfactory work by reason of the
large amount of fluosilicate to be washed out with only a limited
amount of water. At the present time the slime is first stirred
up with the ordinary electrolyte several times, and allowed to
settle, before starting to wash with water at all. The Trail
plant produces daily 8 or 10 cu. ft. of anode residue, of which
over 90 per cent, by volume is solution. The evaporation from
the total tank surface of something like 400 sq. ft. is only about
15 cu. ft. daily; so that only a limited amount of wash- water is
to be used — namely, enough to replace the evaporated water,
plus the volume of the slime taken out.

The tanks are made of 2-in. cedar, bolted together and thor-
oughly painted with rubber paint. Any leakage is caught under-
neath on sloping boards. Solution is circulated from one tank
to another by gravity, and is pumped from the lowest to the
highest by means of a wooden pump. The 22 anodes in each
tank together weigh about 3 tons, and dissolve in from 8 to 10
days, two sets of cathodes usually being used with each set of
anodes. While 300-lb. cathodes can be made, the short-circuiting
gets so troublesome with the spacing used that the loss of capacity
is more disadvantageous than the extra work of putting in and
taking out more plates. The lead sheets used for cathodes are
made by depositing about in. metal on paraffined steel sheets

278 LEAD SMELTING AND REFINING

in four of the tanks, which are different from the others only
in being a little deeper.

The anodes may contain any or all of the elements, gold,
silver, copper, tin, antimony, arsenic, bismuth, cadmium, zincr
iron, nickel, cobalt and sulphur. It would be expected that
gold, silver, copper, antimony, arsenic and bismuth, being more
electronegative than lead, would remain in the slime in the
metallic state, with, perhaps, tin, while iron, zinc, nickel and
cobalt would dissolve. It appears that tin stands in the same
relation to lead that nickel does to iron, that is, they have about
the same electromotive forces of solution, with the consequence
that they can behave as one metal and dissolve and deposit
together. Iron, contrary to expectation, dissolves only slightly,
while the slime will carry about 1 per cent, of it. It appears
from this that the iron exists in the lead in the form of matte.
Arsenic, antimony, bismuth and copper have electromotive
forces of solution more than 0.3 volt below that of lead. As
there is no chance that any particle of one of these impurities
will have an electric potential of 0.3 volt above that of the lead
with which it is in metallic contact, there is no chance that they
will be dissolved by the action of the current. The same is even
more certainly true of silver and gold. The behavior of bismuth
is interesting and satisfactory. It is as completely removed by
this process of refining as antimony is. No other process of
refining lead will remove this objectionable impurity so com-
pletely. Tin has been found in the refined lead to the extent
of 0.02 to 0.03 per cent. This we had no difficulty in removing
from the lead by poling before casting. There is always a certain
amount of dross formed in melting down the cathodes; and the
lead oxide of this reacts with the tin in the lead at a comparatively
low temperature.

The extra amount of dross formed in poling is small, and
amounts to less than 1 per cent, of the lead. The dross carries
more antimony and arsenic than the lead, as well as all the tin.
The total amount of dross formed is about 4 per cent. Table I
shows its composition.

The electrolyte takes up no impurities, except, possibly, a
small part of the iron and zinc. Estimating that the anodes
contain 0.01 per cent, of zinc and soluble iron, and that there
are 150 cu. ft. of the solution in the refinery for every ton of

LEAD REFINING 279

lead turned out daily, in one year the 150 cu. ft. will have taken
up 93 Ib. of iron and zinc, or about 1 per cent. These impurities
can accumulate to a much greater extent than this before their
presence will become objectionable. It is possible to purify
the electrolyte in several ways. For example, the lead can be
removed by precipitation with sulphuric acid, and the fluosilicic
acid precipitated with salt as sodium fluosilicate. By distillation
with sulphuric acid the fluosilicic acid could be recovered, this
process, theoretically, requiring but one-third as much sulphuric
acid as the decomposition of fluorspar, in which the fluorine was
originally contained.

The only danger of lead-poisoning to which the workmen
are exposed occurs in melting the lead and casting it. In this
respect the electrolytic process presents a distinct sanitary
advance.

For the treatment of slime, the only method in general use
consists in suspending the slime in a solution capable of dis-
solving the impurities and supplying, by a jet of steam and air
forced into the solution, the air necessary for its reaction with,
and solution of, such an inactive metal as copper. After the
impurities have been mostly dissolved, the slime is filtered off,
dried and melted, under such fluxes as soda, to a dore bullion.

The amount of power required is calculated thus: Five amperes
in 24 hours make 1 Ib. of lead per tank. One ton of lead equals
10,000 ampere-days, and at 0.35 volts per tank, 3500 watt-days,
or 4.7 electric h.p. days. Allowing 10 per cent, loss of efficiency
in the tanks (we always get less lead than the current which is
passing would indicate), and of 8 per cent, loss in the generator,
increases this to about 5.6 h.p. days, and a further allowance
for the electric lights and other applications gives from 7 to
8 h.p. days as about the amount per ton of lead. At $30 per
year, this item of cost is something like 65c. per ton of lead.
So this is an electro-chemical process not especially favored by
water-power.

The cost of labor is not greater than in the zinc-desilveriza-
tion process. A comparison between this process and the Parkes
process, on the assumption that the costs for labor, interest and
general expenses are about equal, shows that about $1 worth of
zinc and a considerable amount of coal and coke have been done
away with, at the expense of power, equal to about 175 h.p.

280

LEAD SMELTING AND REFINING

hours, of the average value of perhaps 65c., and a small amount
of coal for melting the lead hi the electrolytic method.

More important, however, is the greater saving of the metal
values by reason of increased yields of gold, silver, lead, anti-
mony and bismuth, and the freedom of the refined lead from
bismuth.

Tables II, III, and IV show the composition of bullion, slimes
and refined lead.

Tables V, VI, VII, and VIII give the results obtained ex-
perimentally in the laboratory on lots of a few pounds up to a
few hundred pounds. The results in Tables VI and VII were
given me by the companies for which the experiments were
made.

TABLE I.— ANALYSES OF DROSS
For analyses of the lead from which this dross was taken, see Table II

No.

No. IN
TABLE
II.

Cu.

PER CENT.

As.
PER CENT.

SB.
PER CENT.

FE.
PER CENT.

ZN.

1
2

2
3

0.0005
0.0010

0.0003
0.0008

0.0016

0.0107

0.0016
0.0011

none
a

TABLE II.— ANALYSES OF BULLION

No.

FE.
PER -CENT.

«u

fcO

$
1

Ac.
Oz,p.T.

1
2
3
4
5
6
7

0.0075
0.0115
0.0070
0.0165
0.0120
0.0055
0.0380

0.1700
0.1500
0.1600
0.1400
0.1400
0.1300
0.3600

0.5400
0.6100
0.4000
0.7000
0.8700
0.7300
0.4030

0.0118
0.0158
0.0474
0.0236
0.0432
0.0316

0.1460
0.0960
0.1330
0.3120
0.2260
0.1030
tr.

1.0962
1.2014
1.0738
0.8914
0.6082
0.6600
0.7230

0.0085
0.0086
0.0123
0.0151
0.0124
0.0106
0.0180

98.0200
97.9068
98.1665
97.9014
98.0882
98.2693
98.4580

319.7
350.4
313.2
260.0
177.4
192.5
210.9

2.49
2.52
3.6
4.42
3.63
3.10
5.25

TABLE III.— ANALYSES OF SLIMES

FE.

PER CENT.

Cu.
PER CENT.

SB.
PER CENT.

SN.
PER CENT.

As.
PER CENT.

PB.

ZN.

Bi.

1.27
1.12

8.83
22.36

27.10
21.16

12.42
5.40

28.15
23.05

17.05
10.62

none

LEAD REFINING

281

TABLE IV.— ANALYSES OF REFINED LEAD

Cu.
PER CENT.

As.
PER CENT.

SB.
PER CENT.

FE.
PER CENT.

ZN.

PER CENT.

Ni,Co,CD.
PER CENT.

Bx.
PER CENT.

0.0006

0.0008

0.0005

0.0003

0.0002

0.0010

none

00009

0.0001

0.0009

0.0008

0.24

0.0016

0.0017

0.0014

0.47

none

0.0003

0.0060

0.0003

0.22

00020

00010

00046

022

none

0.0004

none

0.0066

0.0013

none

0.0035

0.14

00004

00038

00004

00035

0.25

00005

00052

0.0004

00039

0.28

0.0003

none

00060

0.0003

0.0049

0.43

0.0003

0.0042

0.0013

0.0059

0.32

0.0005

0.0055

0.0009

0.0049

0.22

0.0005

0.0055

0.0007

0.0091

0.11

0.0004

0.0063

0.0005

0.0012

0.14

1s)

00003

00072

00003

00024

024

0.0006

00062

00012

00083

0.22

0.0006

00072

0.0011

00080

0.23

0.0006

0.0057

00010

0.0053

0.34

0.0005

0,0066

0.0016

0.0140

0.38

0.0005

0.0044

0.0011

0.0108

0.35

*>0

00004

00047

00015

00072

022

0.0004

0.0034

Q.0016

trace

0.23

0.0022

0.0010

0.0046

none

0.0081

0.38

none

TABLE V.— ANALYSES OF BULLION AND REFINED LEAD

Ac.
PER CENT.

Cu.

PER CENT.

SB.
PER CENT.

PB.

PER CENT.

Bullion ....

050

031

0.43

9876

Refined lead

0.0003

0.0007

0.0019

99 9971

TABLE VI.— ANALYSES OF BULLION AND KEFINED LEAD

Cu.

PER CT.

Bi.
PER CT.

As.
PER CT.

SB.
PER CT.

Ac.

OZ.P.T.

Ac.
PER CT.

Au.

OZ.P.T.

FE.
PER CT.

ZN.

PER CT.

Bullion.. .

0.75
0.0027

1.22
.0037

0.936
0.0025

0.6832
0.0000

358.89

1 71

Refined lead .

0.0010

none

0.0022

0.0018

282

LEAD SMELTING AND REFINING

TABLE VII.— ANALYSES OF BULLION, REFINED LEAD AND

SLIMES

6rj
«
w

** S

N*g

W^ £4

Bullion

9673

0096

085

1 42

about 275 1

Refined lead .

00013

000506

0.0028

000068

6 0027

trace

Slimes (dry

sample) ....

9.05

1.9

9.14

29.51

9366.9

0.49

trace

TABLE VIII.— ANALYSES OF BULLION, REFINED LEAD AND

SLIMES

PB.

Cu.

Bi.

Ac.

SB.

As.

PER CENT.

Bullion

87.14

1.40

0.14

0.64

4.0

7.4

Lead..

0.0010

0.0022

0.0017

trace

Slimes.

10.3

9.3

0.52

4.7

25.32

44.58

1 Silver not given. This was the case, also, with the gold in the bullion.
The slimes contained 0.131 per cent, of gold, or 39.1 oz. per ton.

PART X
SMELTING WORKS AND REFINERIES

THE NEW SMELTER AT EL PASO, TEXAS

(April 19, 1902)

In July, 1901, the El Paso, Texas, plant of the Consolidated
Kansas City Smelting and Refining Company l was almost com-
pletely destroyed by fire. The power plant, blast-furnace build-
ing and blast furnaces were entirely destroyed, and portions of
the other buildings were badly damaged. The flames were
hardly extinguished before steps were taken to construct a new,
modern and enlarged plant on the ruins of the old one, and on
April 15, 1902, nine months after the destruction of the former
plant, the new furnaces were blown in. In rebuilding it was
decided to locate the new power-house at some distance from
the other buildings. The furnaces have all been enlarged, each
of the new lead furnaces (of which there are seven) having about
200 tons daily capacity. These and the three large copper fur-
naces have been located in a new position in order to secure a
larger building territory. The entire plant is modern and up to
date in every particular. One of the interesting features is the
substitution of crude oil as fuel in the boiler and roasting depart-
ments. It is intended to use Beaumont petroleum for the gen-
eration of power and the roasting of the ores instead of wood,
coal or coke, and it is expected that a considerable economy
will be effected by this means.

Power Plant. — The power plant is complete in all respects.
It is a duplicate plant in every sense of the word, so that it will
never be necessary to shut the works down on account of the
failure of any one piece of machinery. There are seven boilers,
having a total of 1250 h.p. The four blowers are unusually
large, having a capacity of 30,000 cu. ft. of free air per minute.
They are direct-connected to three tandem compound condensing
Corliss engines. No belts are used in this plant, except for
driving a small blower of 10,000 cu. ft. capacity, which will act
as a regulator. A large central electric plant has been installed
in the power-house, consisting of two direct-connected, direct-
current generators, mounted on the shafts of two cross-compound
condensing Nordberg-Corliss engines. The current from these

1 A constituent company of the American Smelting and Refining Company.

285

286 LEAD SMELTING AND REFINING

generators is transmitted through the plant, operating sampling
works, briquetting machinery, pumps, hoists, motors, cars, etc.,
displacing all the small steam engines and steam pumps used in
the old plant. The power plant is provided with two systems
for condensing; one being a large Wheeler surface condenser, the
other a Worthington central-elevated jet condenser, the idea
being to use the surface condenser during a short period of the
year when the water is so bad that it cannot be used in the boilers.
During the remainder of the year the jet condenser is in service
and the surface condenser can be cleaned. The condensed steam
from the surface condenser, with the necessary additional water,
goes back directly to the boilers when the surface condenser is
in use. The power-house is absolutely fireproof throughout,
being of steel and brick with iron and cement floors. It is pro-
vided with a traveling crane, and no expense has been spared to
make this, as all other parts of the plant, complete in every
respect. The main conductors from the generators pass out
through a tunnel into a brick and steel lightning-arrester house,
from which point the various distributing lines go to different
parts of the plant.

Blast Furnaces. — There are seven large lead furnaces, each
having a capacity of 200 to 250 tons of charge per day, and
three large copper furnaces, each having a capacity of 250 to
300 tons per day. All of the furnaces are enclosed in one steel
fireproof building, the lead furnaces being at one end and the
copper furnaces at the other. Each of the furnaces has its
independent flue system and stack. An entirely new system of
feeding these furnaces has been devised, consisting of a 6-ton
charge car operated by means of a street railroad motor and
controller with third-rail system. The charge cars collect their
charge at the ore beds, limerock and coke storage, and are run
on to 15-ton hydraulic elevators. They are then elevated 38 ft.
to the top of the furnaces, traveling over them to the charging
doors, through which the loads are dumped directly into the
furnaces. This system permits of two men handling about 1000
tons per day. The same system and cars are used for charging
the copper furnaces, except that, as these furnaces are much
lower than the lead furnaces, the charge is dropped into a large
hopper, from which it is fed to the copper furnaces by a man
on the copper furnace 'feed-floor level.

NEW PLANT OF THE AMERICAN SMELTING AND REFINING
COMPANY AT MURRAY, UTAH

BY WALTER RENTON INGALLS

(June 28, 1902)

Murray is a few miles south of Salt Lake City, with which
it is connected by a trolley line. The new works are situated
within a few hundred yards of the terminus of the latter and in
close juxtaposition to the old Germania plant, which is the only
one of the Salt Lake lead-smelting works in operation at present.
The new plant is of special interest inasmuch as it is the latest
construction for silver-lead smelting in the United States, and
may be considered as embodying the best experience in that
industry, the designers having had access to the results attained
a,t almost all of the previous installations. It will be perceived,
however, that there has been no radical departure in the methods,
and the novelties are rather in details than in the general scheme.

The new works are built on level ground; there has been no
attempt to seek or utilize a sloping or a terraced surface, save
immediately in front of the blast furnaces, where there is a drop
of several feet from the furnace-house floor to the slag-yard level,
affording room for the large matte settling-boxes to stand under
the slag spouts. A lower terrace beyond the slag yard furnishes
convenient dumping ground. Otherwise the elevations required
in the works are secured by mechanical lifts, the ore, fluxes and
coal being brought in almost entirely by means of inclines and
trestles.

The plant consists essentially of two parts, the roasting
department and the smelting department. The former com-
prises a crushing mill and two furnace-houses, one equipped with
Bruckner furnaces and the other with hand-raked reverbera-
tories. The reverberatories are of the standard design, but are
noteworthy for the excellence of their construction. Similar
praise may be, indeed, extended to almost all the other parts of
the works, in which obviously no expense has been spared on

287

288 LEAD SMELTING AND REFINING

false grounds of economy. The roasting furnaces stand in a
long steel house; they are set at right angles to the longer axis
of the building, in the usual manner. At their feed end they
communicate with a large dust-settling flue, which leads to the
main chimney of the works. The ore is brought in on a tramway
over the furnaces and is charged into the furnaces through hop-
pers. The furnaces have roasting hearths only. The fire-boxes
are arranged with step-grates and closed ash-pits, being fed
through hoppers at the end of the furnace. The coal is dumped
close at hand from the railway cars, which are switched in on a
trestle parallel with the side of the building, which side is not
closed in. This, together with a large opening in the roof for
the whole length of the building, affords good light and ventila-
tion. The floor of the house is concrete. The roasted ore is
dropped into cars, which run on a sunken tramway passing
under the furnaces. At the end of this tramway there is an incline
up which the cars are drawn and afterward dumped into brick
bins. From the latter it is spouted into standard-gage railway
cars, by which it is taken to the smelting department. The
roasted ore from the Bruckner furnaces is handled in a similar
manner. The delivery of the coal and ore to the Briickners and
the general installation of the latter are analogous to the methods
employed in connection with the reverberatories.

The central feature of the smelting department is the blast-
furnace house, which comprises eight furnaces, each 48 by 160
in. at the tuyeres. In their general design they are similar to
those at the Arkansas Valley works at Leadville. There are
10 tuyeres per side, a tuyere passing through the middle of each
jacket, the latter being of cast iron and 16 in. in width; their
length is 6 ft., which is rather extraordinary. The furnaces
are very high and are arranged for mechanical charging, a rect-
angular brick down-take leading to the dust chamber, which
extends behind the furnace-house. The furnace-house is erected
entirely of steel, the upper floor being iron plates laid on steel
I-beams, while the upper terrace of the lower floor is also laid
with iron plates. As previously remarked, the lower floor drops
down a step in front of the furnaces, but there is an extension
on each side of every furnace, which affords the necessary access
to the tap-hole. The hight of the latter above the lower terrace
leaves room for the large matte settling-boxes, and the matte

SMELTING WORKS AND REFINERIES 289

tapped from the latter runs into pots on the ground level, dis-
pensing with the inconvenient pits that are to be seen at some
of the older works. The construction of the blast furnaces,
which were built by the Denver Engineering Works Company,
is admirable in all respects. The eight furnaces stand in a row,
about 30 ft. apart, center to center. The main air and water
pipes are strung along behind the furnaces. The slag from the
matte-settling boxes overflows into single-bowl Nesmith pots,
which are to be handled by means of small locomotives. The
foul slag is returned by means of a continuous pan-conveyor to
a brick-lined, cylindrical steel tank behind the furnace-house,
whence it is drawn off through chutes, as required for recharging.

The charges are made up on the ground level, immediately
behind the furnace-house. The ore and flux are brought in on
trestles, whence the ore is unloaded into beds and the flux into
elevated bins. These are all in the open, there being only two
small sheds where the charges are made up and dumped into
the cars which go to the furnaces. There are two inclines to the
latter. At the top of the inclines the cars are landed on a trans-
ferring carriage by which they can be moved to any furnace of
the series.

The dust-flue extending behind the furnace-house is arranged
to discharge into cars on a tramway in the cut below the ground
level. This flue, which is of brick, connects with the main flues
leading to the chimney. The main flues are built of concrete,
laid on a steel frame in the usual manner, and are very large.
For a certain distance they are installed in triplicate; then they
make a turn approximately at right angles and two flues continue
to the chimney. At the proper points there are large dampers
of steel plate, pivoted vertically, for the purpose of cutting out
such section of flue as it may be desired to clean. Each flue has
openings, ordinarily closed by steel doors, which give access to
the interior. The flues are simple tunnels, without drift- walls
or any other interruption than the arched passages which extend
transversely through them at certain places. The chimney is of
brick, circular in section, 20 ft. in diameter and 225 ft. high.
This is the only chimney of the works save those of the boiler-
house.

The boiler-house is equipped with eight internally fired corru-
gated fire-box boilers. They are arranged in two rows, face to

290 LEAD SMELTING AND REFINING

face. Between the rows there is an overhead coal bin, from
which the coal is drawn directly to the hoppers of the American
stokers, with which the boilers are provided. Adjoining the
boiler-house is the engine-house; the latter is a brick building,
very commodious, light and airy. It contains two cross-com-
pound, horizontal Allis-Chalmers (Dickson) blowing engines for
the blast furnaces, and two direct-connected electrical generating
sets for the development of the power required in various parts
of the works. A traveling crane, built by the Whiting Foundry
Equipment Company, spans the engine-house. In close prox-
imity to the engine-house there is a well-equipped machine shop.
Other important buildings are the sampling mill and the flue-dust
briquet ting mill.

A noteworthy feature of the new plant is the concrete paving,
laid on a bed of broken slag, which is used liberally about the
ore-yard and in other places where tramming is to be done.
The roast ing-furnace houses are floored with the same material,
which not only gives an admirably smooth surface, but also is
durable. The whole plant is well laid out with service tramways
and standard-gage spur tracks; the intention has been, obviously ,
to save manual labor as much as possible.

THE MURRAY SMELTER, UTAH1

BY O. PUFAHL

(May 26, 1906)

This plant has been in operation since June, 1902. It gives
employment to 800 men. The monthly production consists of
about 4000 tons of work-lead and 700 tons of lead-copper matte
(12 per cent, lead, 45 per cent, copper). The work-lead is sent
to the refinery at Omaha; the matte to Pueblo, Colo. Most of
the ores come from Utah; but in addition some richer lead ores
are obtained from Idaho, and some gold-bearing ores from Nevada.

For sampling the Vezin apparatus is used, cutting out one-
fifth in each of three passes, crushing intervening, the sample
from the third machine being 1-625 of the original ore; after
further comminution of sample in a coffee-mill grinder, it is cut
down further by hand, using a riffle. The final sample is bucked
down to pass an 80-mesh sieve, but gold ores are put through a
120-mesh.

The steps in the smelting process are as follows: Roasting the
poorer ores in reverberatory furnaces and in Bruckner cylinders.
Smelting raw and roasted ores, mixed, in water- jacketed blast
furnaces, for work-lead and lead-copper matte, the latter con-
taining 15 per cent, lead and 10 to 12 per cent, copper. Roasting
the ground matte, containing 22 per cent, of sulphur, down to
J per cent, in reverberatory furnaces. Smelting the roasted
matte together with acid flux in the blast furnace for a matte
with 45 per cent, copper and 12 per cent. lead.

Only the pyritic ores are roasted in Bruckner furnaces, the
lead ores and matte being roasted in reverberatory furnaces.
Each of the 20 Bruckner furnaces, which constitute one battery,
roasts 8 to 12 tons of ore in 24 hours down to 5 J to 6 per cent,
sulphur, with a coal consumption of two tons. The charge weighs
24 tons. The furnaces make one turn in 40 minutes. To increase

1 Translated from Zett. /. Berg.- Hutten.- und Salinenwesen im preuss.
Staate, 1905, LIII, p. 433.

291

292 LEAD SMELTING AND REFINING

the draft and the output, steam at 40 Ib. pressure is blown in
through a pipe; this has, however, resulted in increasing the
quantity of flue dust to 10 to 15 per cent, of the ore charged.
Ten furnaces are attended by one workman with one assistant,
working in eight-hour shifts. For firing and withdrawing the
charge five men are required.

The gases from the Briickners and reverberatory furnaces
pass into a dust-flue 14 x 14 ft. in section and 600 ft. long, built
of brickwork, with concrete vault; in the stack (225 ft. high,
20 ft. diameter) they unite with the shaft-furnace gases, the
temperature of which is only 60 deg.

There are 12 reverberatory furnaces with hearths 60 ft. long
and 16 ft. broad. They roast 14 tons of ore (or 13 tons of matte)
in 24 hours down to 3J to 4 per cent, sulphur, consuming 32 to 34
per cent, of coal figured on the weight of the charge. There are
12 working doors on each side. The small coal (from Rock
Springs, Wyoming), which is burnt on flat grates, contains 5 per
cent, ash and 3 to 5 per cent, moisture. The roasted product is
dumped through an opening in the hearth, ordinarily kept closed
with an iron plate, into cars which are raised by electricity on a
self-acting inclined plane. Their content is then tipped over into
a chute and cooled by sprinkling with water. From here the
roasted matte is conveyed to the blast furnace in 30-ton cars.
The roasted ore is tipped into the ore-bins.

There are eight blast furnaces, 48 x 160 in. at the tuyeres,
of which there are 10 on each of the long sides. The hight
from the tuyeres to the gas outlet is 20 ft., thence to the throat
6 ft.; the distance of the tuyeres from the floor is 4 ft. The base
is water-cooled. The water-jackets of the furnace are 6 ft. high.
The tuyeres (4-in.) are provided with the Eilers automatic ar-
rangement for preventing the furnace gases entering the blast
pipes. The blast pressure is 34 oz. The furnaces are furnished
with the Arents lead wells; the crucible holds about 30 tons of
lead. The slag and the matte run into a brick-lined forehearth
(8 x 3 ft., 4 ft. deep), from which the slag flows into pots holding
30 cu. ft., while the matte is tapped off into flat round pans
mounted on wheels.

The charge is conveyed to the feed-floor by electricity. The
furnace charge is 8000 Ib. and 12 per cent, coke, with 30 per
cent, (figured on the weight of the charge) of " shells" (slag).

SMELTING WORKS AND REFINERIES 293

Occasionally as much as 230 tons of the (moist) charge, exclusive
of coke and slag, has been handled by one furnace in 24 hours.
During one month (September, 1904) 40,000 tons of charge were
worked up, corresponding to a daily average of 166 tons per
furnace.

The lead in the charge runs from 13 to 14 per cent, on an
average. The limestone, which is added as flux, is quarried not
far from the works. The coke used is in part a very ordinary
quality from Utah; in part a better quality from the East, with
9 to 10 per cent. ash. The matte amounts to 10 per cent. The
slag contains 0.6 to 0.7 per cent, lead and 0.1 to 0.15 per cent,
copper. The slag has approximately the following composition:
36 per cent, silica, 23 per cent, iron (corresponding to 29.57 per
cent. FeO), 23 per cent, lime, 3.8 per cent, zinc and 4 per cent,
alumina.

The work-lead is transferred while liquid from the furnaces to
kettles of 30 tons capacity, in which it is skimmed, and thence
cast in molds through a Steitz siphon. First, however, a 5.5-lb.
sample is taken out by means of a special ladle, and is cast into
a plate. From this samples of 0.5 a.t. are punched out at four
points for the assay of the precious metals.

For the purpose of precipitating the flue dust, the blast-
furnace gases are passed into brickwork chambers in which a
plentiful deposition of the heavier particles takes place. From
here the gases go through an L pipe of sheet iron, 18 ft. in diam-
eter, to the Monier flues, which have a cross-section of 256 sq. ft.
and a total length of 2000 ft. A small part of the flues is also
built of brick. The gases unite with the hot roaster gases just
before entering the 225-ft. chimney. In the portion of the blast-
furnace dust first precipitated the silver runs 22 oz. per ton,
while that recovered nearer the stack contains only 8 oz. The
flue dust is briquetted with a small proportion of lime, and, after
drying, is returned to the blast furnaces.

THE PUEBLO LEAD SMELTERS1

t BY O. PUFAHL

(May 12, 1906)

At the Pueblo plant, ores containing over 10 per cent, lead
are not roasted, but are added raw to the charge. For such
material as requires roasting there are in use five Bruckner
furnaces. The charge is 24 tons for 48 to 60 hours; the furnaces
make one revolution per minute and roast the ore down to 6 per
cent, sulphur. There are also two O'Harra furnaces, each roasting
25 tons daily, and 10 reverberatory furnaces 75 ft. in length,
each roasting 15 tons of ore daily down to 4 per cent, sulphur.

The charge for smelting is prepared from roasted ore, together
with Idaho lead ore, Cripple Creek gold ore, briquetted flue dust,
slag and limestone. There are seven water- jacketed furnaces,
which smelt, each, 150 tons of charge per day. The furnaces
have 18 tuyeres, blast pressure 34 oz., cross-section at the tuyeres
48 x 148 in. They are charged mechanically by a car of 4 tons'
capacity.

The output of lead is 11 to 15 tons per furnace. The matte,
which is produced in small quantity, contains 8 to 12 per cent,
lead and the same percentage of copper. It is crushed by rolls,
roasted in reverberatory furnaces, and smelted with ores rich in
silica. The matte resulting at this stage, running 45 to 50 per
cent, in copper, is shipped to be further worked up for blister
copper.

The work-lead is purified by remelting in iron kettles, the
cupriferous dross being pressed dry in a Howard press, and sent
to the blast furnaces. The work-lead is sent to the refineries at
Omaha, Neb., or Perth Amboy, N. J.

To collect the flue dust the waste gases are passed through
long brick flues. The chimneys are 150 to 200 ft. high, and 15 ft.
in diameter. They stand 75 ft. above the ground level of the

1 Abstract from a paper in Zeit. /. Berg.- Hutten.- und Salinenwesen im
preuss. Staate, 1905, LIII, p. 439.

294

SMELTING WORKS AND REFINERIES 295

blast furnaces. The comparatively small proportion of flue dust
produced (0.9 per cent, of the charge) is briquetted, together
with fine ore and 5 per cent, of a thick paste of lime. For this
purpose a White press is used, which makes six briquets at a
time, and handles 10 tons per hour.

According to a tabulation of the results of five months' run-
ning, the proportion of flue dust at several works of the American
Smelting and Refining Company is as follows:

Globe Plant, Denver 0.5 % of the charge.

Pueblo Plant, Pueblo 0.9 % " " "

fillers' Plant, Pueblo 0.5 % " "

East Helena Plant. Helena 0.3 % " " "

Arkansas Valley Plant, LeadviUe 0.2 % " " "

Murray Plant, Murray, Utah 1.2 % " "

The fuel used is of very moderate quality. The coke (from
beehive ovens) carries up to 17 per cent, ash, the coal 10 to 18
per cent. The monthly production is 2300 tons of work-lead
and 150 tons of copper matte (45 to 50 per cent, copper).

At the Eilers plant all sulphide ores, except the rich Idaho
ore, are roasted down to 5 to 7 per cent. S in 15 reverberatory
furnaces, 60 to 70 ft. in length, each furnace roasting 15 tons per
24 hours, in six charges.

The flue dust is briquetted together with fine Cripple Creek
ore, pyrites cinder from Argentine, Kan., Creede ores rich in silica
and 10 per cent. lime. The residue from the zinc smeltery (U. S.
Zinc Company), which is brought to this plant (600 tons a month
containing nearly 10 per cent, lead), is taken direct to the blast
furnaces. Of the latter there are six, each with 18 tuyeres,
which handle per 24 hours 160 to 180 tons of charge, containing
on an average 10 per cent, of lead in the ore, with 10 per cent, of
coke, figured on the charge. The average monthly production
of a furnace is about 360 tons of work-lead, which is purified at
the Pueblo plant. The furnaces are charged by hand. Of the
slag, 30 per cent., as shells, etc., is returned to the charge. The
monthly production of work-lead is 2000 tons, carrying 150 oz.
of silver and 2 to 6 oz. of gold per ton.

The matte amounts to about 8.3 per cent., and contains
12 per cent, copper. It is concentrated up to 45 per cent. Cu,
which is shipped (150 tons a month) for smelting to blister copper.

THE PERTH AMBOY PLANT OF THE AMERICAN
SMELTING AND REFINING COMPANY1

BY O. PUFAHL
(January 27, 1906)

These works were erected in 1895 by the Guggenheim Smelting
Company. They are situated on Raritan Bay, opposite the
southern point of Staten Island, in a position offering excellent
facilities for transportation by land and by water. The materials
worked up are base lead bullion and crude copper, containing
silver and gold, chiefly drawn from the company's smelteries
in the United States and Mexico. Silver ore is received from
South America. The ores and base metals from Mexico and
South America are brought to Perth Amboy by the company's
steamships (American Smelters Steamship Company).

Ore Smelting. — The silver ore from South America (containing
antimony and much silver, together with galena, iron and copper
pyrites) is crushed by rolls and is roasted down from 26 per cent,
to 3 per cent. S in 11 reverberatory furnaces, 70 ft. long and 15
ft. wide (inside dimensions). It is then mixed with rich galena
from Idaho, pyrites cinder, litharge, copper skimmings, and
residues from the desilverizing process, together with limestone,
and is smelted for work-lead and lead-copper matte in three
water- jacketed furnaces, using 12 per cent, coke, figured on the
ore in the charge. Of these furnaces one has 12 tuyeres; it
measures 42 x 96 in. in cross-section at the tuyeres, and 6 ft.
3 in. by 8 ft. at the charging level. The hight of charge is 16 ft.
The other two furnaces have 16 tuyeres each, their cross-section
at the tuyeres being 44 in. by 128 in., at the charging level 6 ft.
6 in. by 12 ft., and hight of charge 16 ft. The furnaces are
operated at a blast pressure of 35 oz. per square inch. The
temperature of the gases at the throat is 140 deg. F. (60 deg. C.)
measured with a Columbia recording thermometer, which works

1 Translated from Zeit. f. Berg.- Hutten.- und Salinenwesen im preuss.
Staate, 1905, LIII, 490.

296

SMELTING WORKS AND REFINERIES 297

very well. These furnaces reduce, respectively, 100 to 120 and
130 to 140 tons of charge per 24 hours; they are also used for
concentrating roasted matte.

Copper Refining. — The crude copper is melted in two furnaces
of 125 tons aggregate daily capacity, and is molded into anodes
by Walker casting machines. Twenty-six anodes are lifted out
of the cooling vessel at a time, and are taken to the electrolytic
plant.

The electrolytic plant comprises two systems, each of 408 vats.
The current is furnished by two dynamos, each giving 4700
amperes at 105 volts. The cathodes remain in the bath for 14
days. The weight of the residual anodes is 15 per cent.

The anode mud is swilled down into reservoirs in the cellar
as at Chrome (De Lamar Copper Refining Company), is cleaned,
dried and refined in a similar manner.

For melting the cathodes there are two reverberatory furnaces
of capacity for 75 tons per 24 hours. The wire-bars and ingots
are cast with a Walker machine. About 3200 tons of refined
copper are produced per month.

Copper Sulphate Manufacture. — The lyes withdrawn from
the electrolytic process are worked up into copper sulphate, shot
copper being added. This latter is prepared in a reverberatory
furnace from matte obtained as a by-product in working up the
lead. About 200 tons of copper sulphate are thus produced per
month; the process used is the same as at the Oker works, Lower
Harz, Germany. The crystals are rinsed, dried and packed in
strong wooden barrels.

Lead Refining. — The working up of the Mexican raw lead is
carried out under the supervision of the customs officers. The
lead, which is imported duty free, must be exported again. From
each bar a sample is cut from above and below by means of a
punch entering half way into the bar. For refining the lead there
are four reverberatory furnaces of 60 tons capacity, with hearths
17 ft. 9 in. by 12 ft. 6 in., a mean depth of 14 in., and a grate
area of 2 ft. 6 in. by 6 ft.; in addition to these there is a furnace
of 80 tons capacity with a hearth 19 ft. 7J in. by 9 ft. 6 in., a
mean depth of 18 in., and grate area of 3 ft. by 6 ft.

For desilverizing the softened lead there are five kettles, each
of 60 tons capacity, 10 ft. 3 in. diameter and 39 in. depth. The
zinc is stirred in with a Howard mechanical stirrer and the zinc

298 LEAD SMELTING AND REFINING

scum is pressed dry in a Howard press, which gives a very dry
scum. The latter is then, while still warm, readily hammered
into pieces for the retorts.

The desilverized lead is refined in five reverberatory furnaces,
of which four take a charge of 50 tons each, and one of 65
tons. The production of desilverized lead is 5000 to 5500 tons
a month.

The distillation of the zinc crusts is carried out in 18 oil-fired
Faber du Faur tilting furnaces. Each retort receives a charge
of 1200 Ib. of broken-up crust and a little charcoal. The distilla-
tion lasts 6 to 7 hours. Fifty gallons of petroleum residues are
consumed per charge. The oil is blown into the furnace with a
compressed air atomizer. After withdrawing the condenser,
which runs on a traveling support, the argentiferous lead is
poured directly from the tilted retort into an English cupel fur-
nace. Seven such furnaces (magnesia-lined, with movable test)
are in use, of which each works up 4.5 to 5 tons of retort metal
in 24 hours. The furnaces are water-jacketed. The blast is
introduced by the aid of a jet of steam. Three tons of coal are
used per 24 hours.

Gold and Silver Parting. — The dore bars are cast into anodes
for electrolytic parting by the Moebius process. The plant con-
sists of 144 cells in 24 divisions. The mean composition of the
electrolytic bath is said to be as follows: 10 per cent, free nitric
acid, 17 grams silver, and 35 to 40 grams copper per liter. The
current is furnished by a 62 k.w. dynamo. One cell consumes
260 amp. at 1.75 volts. One k.w. gives a yield of 1600 oz. fine
silver per 24 hours. The daily production of silver is almost
100,000 oz., and is exceeded at no other works. About $3,000,000
worth of metal is always on hand in the different departments.

THE NATIONAL PLANT OF THE AMERICAN SMELTING
AND REFINING COMPANY1

BY O. PUFAHL
(April 14, 1906)

This plant, at South Chicago, 111., refines base lead bullion.
It comprises four reverberatory furnaces, of which one takes a
charge of 100 tons, one 80 tons, and the other two 30 tons each;
one of the small furnaces is being torn down, and a 120-ton
furnace is to be built in its place. The furnaces are fired with
coal from Southern Illinois, which contains 11 per cent, of ash.

In softening the bullion, the time for each charge is 10 hours.
The first portion tapped consists of dross rich in copper, which
is followed by antimonial skimmings and litharge.

The copper dross is melted up in a small reverberatory fur-
nace, together with galena from Wisconsin (containing 80 per
cent, lead), for work-lead and lead-copper matte, the latter con-
taining about 35 per cent, of copper; this matte is enriched
to 55 per cent, copper by the addition of roasted matte,
and is finally worked up for crude copper (95 per cent.) in a
reverberatory furnace. All the copper so produced is used in
the parting process for precipitating the silver. The antimonial
skimmings are smelted in a reverberatory furnace, together with
coke cinder, for lead and a slag rich in antimony, which is reduced
to hard lead (27 per cent, antimony, 0.5 per cent, copper, 0.5
per cent, arsenic) in a small blast furnace, 14 ft. high, which has
8 tuyeres.

The softened lead is tapped off into cast-iron desilverizing
pots, which usually outlive 200 charges; in isolated cases as many
as 300. For desilverizing, zinc from Pueblo, Colo., is added in
two instalments, being mixed in by means of a Howard stirrer.
After the first addition there remains in the lead 7 oz. of silver
per ton; after the second only 0.2 oz. The first scum is pressed

1 Abstract from a paper in Zeii. /. Berg.- Hutten.- und Salinemoesen im
preuss. Staate, 1905, p. 400.

299

300 LEAD SMELTING AND REFINING

in a Howard press and distilled; the second is ladled off and is
added to the next charge. The Howard stirrer is driven by a
small steam engine suspended over the kettle; the Howard press
by compressed air.

For distilling zinc scum, 12 Faber du Faur tilting retorts,
heated with petroleum residue, are used. The argentiferous lead
(with 9.6 per cent, silver) is transferred from the retort to a pan
lined with refractory brick, which is wheeled to the cupelling
hearth and raised by means of compressed-air cylinders, so as to
empty its molten contents through a short gutter upon the cupel-
ling hearth. The cupelling hearths are of the water-cooled English
type, and are heated by coal with under-grate blast. The cast-
iron test rings, with reinforcing ribs, are made in two pieces,
slightly arched and water-cooled; they are rectangular, with
rounded corners, and are mounted on wheels. The material of
the hearth is marl.

Argentiferous lead is added as the operation proceeds, and
finally the dore bullion is poured from the tilted test into thick
bars (1100 oz.) for parting.

The desilverized lead is refined in charges of 28 tons (4 to 5
hours) and 80 to 90 tons (8 to 10 hours), introducing steam through
four to eight half-inch iron pipes. The first skimmings contain a
considerable proportion of antimony and are therefore added to
the charge when reducing the antimonial slags in the blast furnace.
The litharge is worked up in a reverberatory furnace for lead of
second quality. The refined lead is tapped off into a kettle,
from which it is cast into bars through a siphon.

The parting of the dore" bullion is carried out in tanks of gray
cast iron, in which the solution is effected with sulphuric acid of
60 deg. B. The acid of 40 deg. B. condensed from the vapors is
brought up to strength in leaden pans. In a second larger tank,
which is slightly warmed, a little gold deposits from the acid
solution of sulphates. The solution is then transferred (by the
aid of compressed air) to the large precipitating tank, and diluted
with water. It is here heated with steam, and the silver is
rapidly precipitated by copper plates (125 plates 18 x 8 x 1 in.)
suspended in the solution from iron hooks covered with hard
lead. After the precipitation, the vitriol lye is siphoned off, the
silver is washed in a vat provided with a false bottom, is removed
with a wooden shovel, and is pressed into cakes 10 x 10 x 6 in.

SMELTING WORKS AND REFINERIES 301

The refining is finished on a cupelling hearth fired with petro-
leum residue, adding saltpeter, and removing the slag by means
of powdered brick. After drawing the last portion of slag the
silver (0.999 fine) is kept fused under a layer of wood-charcoal for
20 minutes, and is then cast into iron molds, previously blackened
with a petroleum flame. The bars weigh about 1100 oz.

The gold is boiled with several fresh portions of acid, is washed
and dried, and finally melted up with a little soda in a graphite
crucible. It is 0.995 fine.

The lye from the silver precipitation, after clearing, is evapo-
rated down to 40 deg. B. in leaden pans by means of steam coils,
and is transferred to crystallizing vats. The first product is-
dissolved in water, the solution is brought up to 40 deg. B. strength,
and is allowed to crystallize. The purer crystals so obtained are
crushed, and are washed and dried in centrifugal apparatus;
they are then sifted and packed in wooden casks in two grades
according to the size of grain. The very fine material goes back
into the vats. From the first strongly acid mother liquor, acid
of 60 deg. B. is prepared by concentrating in leaden pans, and
this is used for the parting operation.

THE EAST HELENA PLANT OF THE AMERICAN SMELTING
AND REFINING COMPANY1

BY O. PUFAHL

(April 28, 1906)

The monthly production of these works is about 1500 tons of
base bullion (containing 150 oz. Ag and 4 to 6 oz. Au per ton),
and 200 tons of 45 per cent, copper matte. The base bullion is
shipped to South Chicago, the matte to Pueblo.

The ore-roasting is done in two batteries of eight reverberatory
furnaces and 16 Bruckner furnaces, the resulting product con-
taining on an average 20 per cent, lead and 3 per cent, sulphur.
The charge for the blast furnaces consists of roasted ore, rich
galena, argentiferous red hematite, briquetted flue dust, slag
(shells) from the furnace itself, lead skimmings, scrap iron and
limestone.

Four tons of the charge are dumped over a roller into a low
car, which is then drawn up an inclined plane to the charging
gallery by an electric motor and is then dumped into the furnace.

The two rectangular blast furnaces (Eilers* type) have eight
tuyeres on each of their longer sides and cast-iron water-jackets
of 6 ft. hight. The blast is delivered at a pressure of 40 oz.
The lead is drawn off through a siphon tap into a cooling kettle.
The furnace has a large forehearth for separating the matte and
the slag. The slag is received by a two-pot Nesmith truck,
having an aggregate capacity of 14 cu. ft. These trucks are
hauled to the dump by an electric locomotive. The shells are
returned to the furnace with the charge.

The matte (with about 6 per cent. Cu and the same percentage
of lead) is tapped off into iron molds and after cooling is crushed
to 0.25-in. size, to be roasted in the reverberatory furnaces and
smelted up together with roasted ore for a 15 per cent, matte.
The latter is crushed, roasted and separately smelted together

1 Abstract from a paper in Zeit. /. Berg.- Hiitten- und Salinenweaen im,
preuss. Staate, 1905, p. 400.

302

SMELTING WORKS AND REFINERIES 303

with silicious ore for 45 per cent, matte, which is then sent to
Pueblo to be worked up into blister copper. The small quantity
of speiss which is formed is broken up and returned to the blast
furnaces with the charge. The slag contains 0.5 to 0.8 per cent,
lead and 0.5 oz. silver per ton.

THE GLOBE PLANT OF THE AMERICAN SMELTING AND
REFINING COMPANY1

BY O. PUFAHL
(May 5, 1905)

This plant produces 1800 tons of base bullion per month and
200 tons of lead-copper matte containing 45 to 52 per cent, of
copper. The ores smelted are mostly from Colorado, but include
also galena from the Coeur d'Alene and other supplies. The
limestone is quarried 14 miles from Denver; coke and coal are
brought from Trinidad, Colo.

All sulphides, except the slimes, concentrates and the rich
Idaho ores, are roasted. For roasting there are:

(1) Fifteen reverberatory furnaces, five of which measure
60 x 14 ft., and the other ten 80 x 16 ft. externally. In 24 hours
these roast six charges of 4400 Ib. (average) of moist ore (2.15
tons of dry ore) from 28 to 30 per cent, down to 3 to 4 per cent,
sulphur. Each furnace is attended by three men working in
12-hour shifts; the stoker earns $2.75; the roasters, $2.30.

(2) Two Brown-O'Harra furnaces, 90 ft. long, with two
hearths, and a small sintering furnace attached. They have
three grates on each long side, and each roasts 26 tons of ore in
24 hours down to J per cent, sulphur.

(3) Twelve Bruckner furnaces, each taking 24 tons' charge,
with under-grate blast, the air being fed into the cylinders by a
steam jet. According to the zinc content of the ores the roasting
operation lasts 70 to 90 hours, the furnace making one revolution
per hour. The roasted product from the Bruckner furnaces is
pressed into briquets, together with fine ore, flue dust and lime.

The smelting is carried out in seven blast furnaces, with
16 tuyeres, blast at 2-lb. pressure, hight of furnace 18 ft. 6 in.,
section at the tuyeres 42 x 144 in. The charge is 120 to 150
tons exclusive of slag and coke. The slag and the matte are

1 Abstract from an article in Zett. /. Berg.- Hutten.- und Salinenwesen im
preuss. Staate, 1905, LIII, p. 444.

304

SMELTING WORKS AND REFINERIES 305

tapped off together into double-bowl Nesmith cars, which are
hauled, by an electric locomotive, to a reverberatory furnace
(hearth 20 x 12 ft.) in which they are kept liquid, for several
hours, in charges of 14 to 15 tons, in order to effect complete
separation. A little work-lead is obtained in this operation,
while the matte is tapped off into cast-iron pans of one ton capac-
ity, and the slag, 0.5 to 0.6 per cent, lead, 0.6 to 0.7 oz. silver, is
removed in 5-ton pots to the dump.

The matte is broken up, crushed to 0.25 in. size, roasted in
the reverberatory furnaces, smelted for a 45 to 52 per cent,
copper matte, which is shipped to be further worked up into
blister copper. The crude matte contains 10 to 12 per cent,
copper, 12 to 15 per cent, lead, 40 oz. silver and 0.05 oz. gold.

From the siphon taps of the blast furnaces the work-lead is
transferred to a cast-iron kettle of 33 tons' capacity. Here the
copper dross is removed, the metal is mixed by introducing
steam for 10 minutes, sampled, and the lead is cast into bars
through siphons. It contains about 2 per cent, antimony, 200 oz.
silver and 8 oz. gold. This product is refined at Omaha.

The blast-furnace gases pass through a flue 1200 ft. long, and
enter the bag-house, in which they are filtered through 4000
cotton bags 30 ft. long and 18 in. in diameter. These bags are
shaken every 6 hours. The material which falls to the floor is
burnt where it lies, sintered and returned to the blast furnaces.

In the engine house there are four Connersville blowers, two
of which are No. 8 and two of No. 7 size. Each blast furnace
requires 45,000 cu. ft. of air a minute.

The works give employment to 450 men, whose wages (for
10- to 12-hour shifts) are $2 to $3.

LEAD SMELTING IN SPAIN

BY HJALMAR ERIKSSON

(November 14, 1903)

A few notes, gathered during a couple of years while I was
employed at one of the large lead works in the southeastern part
of Spain, are of interest, not as showing good work, but for com-
paring the results obtained in modern practice with those obtained
by what is probably the most primitive kind of smelting to be
found today. The plant about to be described may serve as a
general type for that country. As far as I know, the exceptions
are a large plant at Mazarron, fully up to date and equipped with
the most modern improvements in every line; a smaller plant at
Almeria, also in good shape, and the reverberatory smelting of
the carbonates at Linares. It should be kept in mind, however,
that the conditions are peculiar, iron and machinery being very
expensive and manual labor very cheap.

About 4 ft. above the tuyeres the furnace is built of uncalcined
brick made of a black graphitic clay found in the mines near by;
the upper part is of common red brick. The entire cost of one
furnace does not reach $100. The flue leads to a main gallery
3.5 by 7 ft., which goes down to the ground, and extends several
times around a hill, the chimney being placed on the top of the
hill, considerably above the furnace level. The gallery is about
10,000 ft. long, and is laid down in the earth, with the arched
roof just emerging. It is all built of rough stone, the inside being
plastered with gypsum. The furnace has three tuyeres of 3-in.
diameter. The blast pressure is generally 4 to 6 in. of water.
Neither feeding floor nor elevators are used, only a couple of
scaffolds, the charge being lifted up gradually by hand in small
convenient buckets made of sea-grass. When charging the fur-
nace, coke is piled up in the center, and the mixture of ore, fluxes
and slag is charged around the walls. The slag and matte are
left to run out together on an inclined sand-bed. The matte,
flowing more quickly, goes further and leaves the slag behind,

306

SMELTING WORKS AND REFINERIES

307

but the separation thus obtained is, of course, very unsatisfactory.
The charge mixture is weighed and made for each furnace every
morning. When it is all put through, the furnace is run down
very low, without any protecting cover on the top; several iron
bars are driven through the furnace at the slag-tap level, for

FIG. 41. — Spanish Lead Blast
Furnace.

holding up the charge; the lead is all tapped out; a big hole is
made in the crucible for the purpose of cleaning it out; all accre-
tions are loosened with a bar; the hole is closed with mud of the
graphitic clay; bars are removed, when the crucible is filled with

308 LEAD SMELTING AND REFINING

coke from the center and the charging is continued. In this way
a furnace can be kept running for any length of time, but at a
great loss of heat, and with a great increase of flue dust.

The current practice, in many parts of Spain, is to run the
same number of ore-smelting and of matte-smelting furnaces.
All the slag and the raw matte, produced by the ore-smelting
furnaces, is re-smelted in the matte furnaces, together with some
dry silver ores. No lead at all is produced in the matte furnaces,
only a matte containing up to 150 oz. silver per ton and 25 to 35
per cent, of the lead charged on them. This rich matte is calcined
in kilns, and smelted together with the ore charge.

The ores we smelted were galena ranging from 5 to 83 per
cent, lead and about 250 oz. silver per ton of lead; dry silver ores
containing up to 120 oz. silver per ton, and enough of the Linares
carbonates for keeping the silver below 120 oz. per ton in the
lead. The gangue of the galena was mainly iron carbonate.
Most of that ore was hand picked and of nut size. Machine
concentrates with more than 30 per cent, lead or containing
much pyrite were calcined; everything else was smelted raw.
The flux exclusively used, before I came, was carbonate of iron,
which, by the way, was considered a "cure-for-all." The slag
analyses showed:

CaO, below 4 per cent. A12O3, 5 to 10 per cent.

FeO, above 45 per cent. Pb, by fire assay, 0.75 to 2.5 per cent.

SiO2, about 30 per cent. Ag, by fire assay, 2 to 3 oz. per ton.
BaO, 5 to 10 per cent.

The specific gravity of the slag was about 5, or practically
the same as that of the matte. The output of metallic lead was
about 70 per cent.; of silver, 84 per cent. The working hight
of the furnaces — tuyere level to top of charge — was at that
time only 7 ft., and I was told that it had been still lower
before.

To the working hight of the furnaces was added 2 ft., simply
by putting up the charging doors that much. A very good
limestone was found just outside the fence around the plant.
Enough limestone was substituted for the iron carbonate, to keep
the lime up to 12 per cent, in the slag, reducing the FeO to below
35 per cent, and the specific gravity to below four.

The result of these alterations was an increase in the output

SMELTING WORKS AND REFINERIES 309

of metallic lead, from 76 to 85 per cent.; of silver from 84 to 90
per cent.; a comparatively good separation of slag and matte,
and a slag running about 0.5 to 0.75 per cent. Pb and 1.5 oz. Ag
per ton.

Owing to the great extent of the gallery, and the consequent
good condensation of the flue dust, the total loss of lead and
silver was much smaller than would be expected; in no case being
found above 4 per cent.

The composition of the charge was 55 per cent, ore and roasted
matte, 13 per cent, fluxes, and 32 per cent. slag. Coke used was
11 per cent, on charge, or 20 per cent, on ore smelted. Each
furnace put through 10 to 15 tons of charge, or 7 tons of ore, in
24 hours. Eight men and two boys were required for each
furnace, including slag handling and making up of the charge.
The cost of smelting was 17 pesetas per ton of ore, which at the
usual premium (£1 = 34 pesetas = $4.85) equals $2.43. This
•cost is divided as follows:

Coke $1 .47

Fluxes 0.04

Labor 0.65

Coal for power 0 . 10

General expenses 0 . 17

Total $2.43

This $2.43 per ton includes all expenses of whatever kind.
The iron carbonate flux contained lead and silver, which was not
paid for. The fluxes are credited for the actual value of this
lead and silver. Without making this discount, the cost of flux
would amount to 26c. per ton, making the entire smelting cost
come to $2.65. As an explanation of the low cost of labor,
it may be noted that the wages were, for the furnace-man,
2.25 pesetas, or 32c. a day; for the helpers, 1.75 pesetas, or 25c.
a, day.

The basis for purchasing the galena ore may here be given,
reduced to American money; lead and silver are paid for ac-
cording to the latest quotations for refined metals given by the
Revista Minera, published at Cartagena. (The quotations are
the actual value in Cartagena of the London quotations.)

The following discounts are made: 5 per cent, for both silver

310 LEAD SMELTING AND REFINING

and lead; $6.40 per ton on ore containing 7 per cent. Pb and
below; this rises gradually to a discount of $7.75 per ton of ore
containing 30 per cent. Pb and above.

The transportation is paid by the purchaser and amounts to
about $1.20 per ton of ore.

The dry silver ores were cheaper than this and the lead car-
bonates much more expensive.

LEAD SMELTING AT MONTEPONI, SARDINIA1

BY ERMINIO FERRARIS

(October 28, 1905)

In dressing mixed lead and zinc carbonate ores by the old
method of gradual crushing with rolls, middling products were
obtained, which could be further separated only with much loss.
Inasmuch as the losses in the metallurgical treatment of such
mixed ore were reckoned to be less than in ore dressing, these
between-products at Monteponi were saved for a number of
years, until there should be enough raw material to warrant the
erection of a small lead and zinc smeltery.

In 1894 the lead smeltery in Monteponi was put in operation;
in 1899 the zinc smeltery was started. At about the same time
the reserves of lead ore were exhausted, and the lead plant then
began to treat all the Monteponi ores and a part of those from
neighboring mines.

As will be seen from the plan (Fig. 42), the smelting works
cluster in terraces around the mine shaft, covering an area of
about 3000 sq. m. (0.75 acre); the ore stocks and the pottery of
the zinc works are located in separate buildings.

During the first years of working, the slag had purposely been
kept very rich in zinc, in the hope of utilizing it later for the
production of zinc oxide. It had an average zinc content of
16.80 per cent., or 21 per cent, of zinc oxide, with about 32 per
cent. SiO2, 25 per cent. FeO, and 14 per cent. lime. According
to the recent experiments, this slag can very well be used for
oxide manufacture, in connection with calamine rich in iron.
The slag made at the present time has only 15 per cent. ZnO;
25 per cent. SiO2; 16 per cent. CaO; 3 per cent. MgO; 33 per cent.
FeO; 2.5 per cent. A12O3, and 2 per cent. BaO, and small quantities
of alkalies, sulphur and lead (1 to 1.5 per cent).

The following classes of ore are produced at Monteponi:

1. Lead carbonates, with a little zinc oxide; these ores are

1 Translated from Oest. Zeit. /. Berg.- und Huttenwesen, 1905, p. 455.

311

312

LEAD SMELTING AND REFINING

screened down to 10 mm. The portion held back by the screen
is sent straight to the shaft furnaces; the portion passing through
is either roasted together with lead sulphides, or is sintered by
itself, according to circumstances.

2. Dry lead ores, mostly quartz, with 10 to 15 per cent, lead,
which are mixed for smelting with the lead carbonates.

FIG. 42. - General Plan of Works.

3. Lead sulphides, which are crushed fine and roasted dead.
Quartz sand is added in the roasting, in order to decompose the
lead sulphate and produce a readily fusible silicate; as quartz
flux, fine sand from the dunes on the coast is used. This is a
product of decomposition of trachyte, and contains 88 per cent,
of silica, together with alkalies and alumina. The roast is effected

SMELTING WORKS AND REFINERIES

313

in two hand-raked reverberatory furnaces, 18 m. long, which
turn out 12,000 kg. of roasted ore in 24 hours, consuming 1800 kg.
of English cannel coal, or 2400 kg. of Sardinian lignite. There
is also a third reverberatory furnace, provided with a fusion
chamber, which is used for roasting matte and for liquating
various secondary products.

The charge for the shaft furnace, as a rule, consists of 50 per
cent, ore (crude and roasted), 20 per cent, fluxes and 30 per cent.
slag of suitable origin. The fluxes used are limestone from the
mine, containing 98 per cent. CaCO3, and limonite from the
calamine deposits. This iron ore contains 48 per cent. Fe, not
more than 4 per cent. Zn, a little lead and traces of copper and
silver.

A shaft furnace will work up a charge of 60 tons, equal to
30 tons of ore, in 24 hours, with a coke consumption of 12 per

FIG. 43. — Elevation of works on line A B C D E F of Fig. 42.

cent, of the weight of the charge and a blast pressure of 50 mm.
of mercury. There are three furnaces, of which two are used
alternately for smelting lead ores, while one smaller furnace serves
for smelting down products, such as hard lead, copper matte and
copper bottoms.

Figs. 43 to 46 show one of the furnaces. It will be seen at
once that its construction is similar to that of the standard
American furnaces. Pilz furnaces were tried in the first few years,
but were finally abandoned, as they could not be kept running
for any satisfactory length of time with slags rich in zinc. Dilut-
ing the slag, on the other hand, would have led to an increased
coke consumption, and would have rendered the slag itself worth-
less. The furnace, however, differs in several respects from its
American prototype; the following are some of the chief charac-
teristics peculiar to it:

314

LEAD SMELTING AND REFINING

The chimney above the feed-floor covers one-third of the
furnace shaft, and is turned down in the form of a siphon, to
connect with the flue-dust chamber. The lateral faces, which
are left open, serve as charging apertures; the central one of
these, provided with a counterbalanced sheet-iron door, is used
for charging from cars. The square openings at the ends, which
are covered with cast-iron plates, are used for barring down the
furnace shaft and may also be used for charging. By this ar-
rangement, together with the two hoppers placed laterally on
the chimney, it is possible to distribute the charge in any desired

Section E F; Section G H.

FIG. 44. — Shaft Furnace for Lead Smelting.

manner over the whole cross-section of the furnace. This arrange-
ment greatly facilitates the removal of any accretions in the
furnace shaft, as the centrally placed chimney catches all the
smoke, while the charge-holes render the furnace accessible on
all sides. In case of large accretions being formed, the whole
furnace can be emptied, cleaned and restarted in 24 to 36 hours.
The smelting cone is enclosed by cast-steel plates 50 cm.
high, instead of having a water-jacket. These are cooled as
desired by turning a jet of water on them. The plates are con-

SMELTING WORKS AND REFINERIES

315

nected to the furnace shaft by a bosh wall 25 cm. thick, which
is surrounded with a boiler-plate jacket. These jacket plates
also are cooled from the outside by sprays of water. With this
arrangement the consumption of water is less than with water-
jackets, as a part of the water is vaporized, and the danger of
leakage of the jackets is avoided. The cast-steel plates are
made in two patterns; there are two similar side-plates, each
with four slits for the tuyeres, and two end-plates, provided with
a circular breast of 30 cm. aperture, for tapping the slag. The
breast is cooled by water flowing down, and is closed in front by
a plate of sheet iron, in which is the tap-hole for running off the

Section J L. Section C D.

FIG. 45. — Shaft Furnace.

slag. When cleaning out, this sheet-iron plate is removed and
the breast is opened, thus providing easy access to the hearth.
The four cast-steel plates are anchored together with bolts at
their outer ribs, and rest on two long, gutter-shaped pieces of
sheet iron, which carry off all the water which flows down, and
keep it away from the brickwork of the hearth.

The hearth, cased with boiler plate and rails, has at the side
a cast-iron pipe of 10 cm. diameter for drawing off the lead to
the outside kettle; this pipe has a slight downward inclination,,
to prevent the slag flowing out; every 20 minutes lead is tapped,,
and the end of the pipe is then plugged up with clay.

316

LEAD SMELTING AND REFINING

The furnace shaft is supported upon a hollow mantel, which
serves at the same time as blast-pipe. The blast-pipe has eight
lateral tees, which are connected by canvas hose with the eight
tuyeres. The mouth of the tuyeres has the form of a horizontal
slit, whereby the air is distributed more evenly over the entire
zone of fusion.

The precipitation of flue dust is effected in a brick condensing
chamber, placed near the beginning of the main flue. The main
flue terminates on the hill (see Fig. 43) in a chimney, the top of
which is 160 m. above the ground level of the works, affording
excellent draft. The condensing chamber (Figs. 49 to 51) con-
sists of a vaulted room, 3.40 m. wide and 6.60 m. long, which is

FIG. 46. — Shaft Furnace for Lead
Smelting. (Section A B.)

divided into twelve compartments by one longitudinal and five
baffle walls. The gases change direction seven times, and pass
over the longitudinal wall six times, being struck six times by
fine sprays of water. The six atomizers for this purpose consume
1.5 liter of water per minute, of which four-fifths is vaporized,
while one-fifth flows off to the lower water basin. By this means
10 to 15 per cent, of the total flue dust is precipitated in the
condensing chamber itself, and is removed from time to time as
mud through the lower openings, which are water-sealed. The
remainder of the volatilized water precipitates the flue dust
almost completely on the way to the stack, so that only a short
column of steam is visible at the mouth of the stack. The flue
to the stack passes for the most part underground through aban-

SMELTING WORKS AND REFINERIES

319

donei adits and galleries, thus providing a variety of changes in
cross-section and in direction, and assisting materially the action
of the condensing chamber.

As the charge of the shaft furnaces is poor in sulphur, no
real matte is produced, but only work lead and lead ashes (Blei-
schaum), which contains 90 per cent, of lead, 1.6 per cent, sulphur,
0.4 per cent, zinc, 0.85 per cent. Cu., 0.99 per cent. Fe, and 0.22
per cent. Sb. By liquation and a reducing smelt in a reverbera-
tory furnace, most of the lead is obtained, along with a lead-
copper matte, which is smelted for copper matte and antimonial
lead in the blast furnace.

PIG. 49. — Fume Condenser. (Section A B.)

The copper matte, containing 18 per cent. Cu, 25 per cent. Fe,
30 per cent. Pb and 18.4 per cent. S, is roasted dead in a rever-
beratory furnace, is sintered, and melted to copper-bottoms in a
small shaft furnace. These copper-bottoms, which contain 60
per cent, copper and 25 per cent, lead, are subjected to liquation,
and finally refined to blister copper.

The zinc-desilvering plant, Fig. 47, consists of a reverberatory
softening furnace, two desilvering kettles of 14 tons capacity,
a pan for liquating the zinc crust, and a small kettle for receiving
the lead from the liquation process.

This pan has the advantage over the ordinary liquating kettle,
that the lead which drips off is immediately removed, before it
can dissolve the alloy; the silver content of the liquated lead is

320

LEAD SMELTING AND REFINING

FIG. 50, - Fume Condenser. (Section E F G H.)

FIG. 51. — Fume Condenser.
(Section C D.)

SMELTING WORKS AND REFINERIES 321

scarcely 0.05 per cent., while the dry alloy contains 5 to 8 per
cent.

The removal of the zinc is effected in a second reverberatory
furnace. Formerly the steam-method was used, but the rapid
wear of the kettles, and the excessive formation of oxides called
for a change in the process. The zinc-silver alloy is distilled in
a crucible of 200 kg. capacity, and is cupeled in an English cupel
furnace. The details of the reverberatory furnace are shown in
Fig. 48.

The composition of the final products is shown by the following
analyses: Lead: Zn, 0.0021 per cent.; Fe, 0.0047 per cent.; Cu,
0.0005 per cent.; Sb, 0.0030 per cent.; Bi, 0.0007 per cent.; Ag,
0.0010 per cent.; Pb, 99.998 per cent. Silver: Ag, 99.720 per
cent.; Cu, 0.121 per cent.; Fe, 0.005 per cent.; Pb, 0.018 per cent.;
Au, 0.003 per cent.

INDEX

PAGE

Alloy, retorting the, in lead re-
fining 267

Alumina, experience with 259

American Smelting and Refining

Co. 4, 6, 26, 93, 113, 252, 295

at Murray, Utah 287

Atmosphere, effect of on con-
crete 242

Bag-house, cost of attending. . . 246

standard 246

Bag-houses for saving fume .... 244

Bartlett, Eyre 0 244

Bayston, W. B 199

Bennett, James C 66

Betts, Anson G 270, 274

Between products, working up

of 39

Biernbaum, A 41, 148, 160

Blast furnace of circular form . . 253

Spanish lead 307

Blast, volume and pressure of hi

lead smelting 76

Blower, rotary, deficiency of . . . 251
Blowers for lead and copper

smelting 256

now more powerful for lead

smelting use 252

Blowers, rotary, method of test-
ing volumetric efficiency

of 254

vs. blowing engines 254

vs. blowing engines for lead

smelting 251

Blowing engines, when to use. . 259

Bonne Terre lead deposits 18

orebody, Missouri 13,14

Borchers, W 114, 116, 127

PAGE

Bormettes method, combination

processes in 222

Bradford, Mr 55

Bretherton, S. E 251, 258

Broken Hill Proprietary Block 14, 59

Broken Hill practice 51

Proprietary Co. 52, 113, 124
145, 175, 178, 206
Bricking plant for flue dust and

fine ores 66-70

Briquetting costs 62

methods of avoiding. . . .63, 64

process, operations in 59

Bullion, analyses of in lead re-
fining 281

refined lead and slimes,

of . . 282

Canadian Smelting Works 275

Carlton Iron Co 63

Carmichael, A. D 56, 199

Carmichael-Bradford process 175-185

brief estimate of 209

claims of in patent 199

recommendations of 124

process, points concerning. 131
Cement walls, how to build. . . . 241

Channing, J. Parke 254

Charge-car in smelting, true

function of 94

feeding of in lead smelting. 77
mechanical character of in

lead smelting 78

Charges, effect of large hi lead

smelting 77

Cherokee Lanyon Smelter Co ... 104

Chimney bases 237

Chisholm, Boyd & White Co. . . 64

323

324

INDEX

PAGE

Clark, Donald 114, 144, 175

Coeur d'Alene mines 5, 6, 7

Concrete flues and stacks, ad-
vantages and disadvan-
tages of 242

in metallurgical construc-
tion 234

Connersville Blower Co 252

Consolidated Kansas City Smelt-
ing and Refining Co 285

Coke, percentage necessary to

use in smelting 259

Croll, H. V 253

Cupellation in lead refining. . . . 269

De Lamar Copper Refining Co. 297
Desilverization in lead refining. . 265
Desloge practice contrasted with

others 46

Doeltz, F. 0 139

Dross, analyses of in lead refin-
ing 279

Dupuis & Sons 63

Dust chamber, arched form .... 231

beehive form of 232

design 229

rectangular form 230

concrete 235-237

Dwight. Arthur S 73, 81

spreader and curtain in fur-
naces 91

East Helena and Pueblo smelt-
ing systems compared ... 93

plant of the American Smelt-
ing and Refining Co. ... 302

system of smelting 88-94

Edwards, Henry W.. . .234, 240, 242

Einstein silver mine 14

Engine, blowing, proper field of 257

blowing, and rotary blowers 258
Eriksson, Hjalmar 306

Federal Lead Co 38

Mining and Smelting Co. . . 7
Feeders, cup and cone, for

round furnaces . . 81

PAGE

Ferraris, Erminio 311

Flat River mines 18

Flue gases and moisture, effect

of on concrete 242

Flues, concrete 234, 240, 242

Foundations for dynamos 236

Fremantle Smelting Works .... 145

Fume-smelting, cost of 33

in the hearth 32

Furnace operations at Desloge,

Mo 45

Furnaces at Desloge, Mo 43

reverberatory, at Desloge,

Mo 42

Galena, experiments hi roasting 129

lime-roasting of 14

new methods of desulphur-
izing 116

roasting of by Savelsberg

process 122, 123

Gas, furnace, effect of on cement 240
Gelatine, use of in electrolytic

lead refining 275

Germot, A 224

process 224

Globe plant of the American

Smelting and Refining Co. 304
Smelting and Refining Co. . . 244

Greenway, T. J 59

Guillemain, C 133

Harvard, Francis T 242

Hearth, covered-in 36

Heat, effect of on cement 242

Heberlein, Ferdinand. .113, 167, 199

Hixon, Hiram W 256, 258

Harwood, E. J 51

Hourwich, Dr. Isaac A 27

Huntington-Heberlein process . .113,

144-147

consideration and estimate
of 203-209

credit due to 126

process as distinguished
from others 118

economic results of ... 155-159

INDEX

325

PAGE

Huntington-Heberlein explained

by the inventors .... 167-173
process at Friedrichshiitte . 148
process, from the hygienic

standpoint 160

ideas of in patent specifica-
tions 117

process, introduction of at

Tarnowitz, Prussia 41

and Savelsberg processes,
essential difference be-
tween 192

process, some disadvantages

of 165, 166

Huppertz, L 121

Hutchings, W. Maynard 108, 126, 170
Huntington, Thomas. . . 113, 167, 199

lies, Malvern W 96, 252

Ingalls, W. R. 3, 16, 27, 42, 177, 186,
193, 215, 224, 244, 287
Iron, behavior of in silver-lead

smelting 75

Jackson Revel mine 14

Johnson, E. M 104

R. D. 0 18

Jones, Richard 244

Samuel T 244

Laur, F 224

Lead, analyses of refined 281

bullion, electrolytic refining

of base 270

bullion, Parkes process of
desilverizing and refining 263

bullion, softening of 263

concentrate Joplin district,

valuation of 25

and copper smelting, the

Bormettes method of 215-223
deposits, southeastern Mis-
souri 18

Joplin district 8

marketing 3

-ore roasting, consideration
of new processes 135-138

PAGE
Lead ore, average prices for. ... 27

ore, cost of smelting 32

-ore roasting, theoretical

aspects of 133

ores, Galena, Kan 24

ores, method of valuing. . . 26
ores, southwestern Missouri 24

Park City, Utah 8

-poisoning in old and new

processes 162-165

refining, electrolytic 274

soft, Missouri 25

smelting at Desloge, Mo. . . 42
smelting at Monteponi, Sar-
dinia 311

smelting and refining, cost

of 96

smelting in the Scotch

hearth 31

smelting in Spain 306

smelting at Tarnowitz, Prus-
sia 41

source of in Missouri 13

in southeastern Missouri 7, 10, 17
sulphide and calcium sul-
phate, metallurgical be-
havior of 139-143

total production United

States 5

yield from Scotch hearths. 39

Leadville, Colo., mines 8

Lewis, G. T 244

Lime-roasting of galena 126

Lotti, Alfredo 215

Messiter, Edwin H 229, 240

Middleton, K. W. M 31

Mine La Motte 14

Minerals, briquetting of 63

Mining methods in Missouri . . . 19-23

Missouri Smelting Co 197

Mould, H. S., Co 64

Murray smelter, Utah 291

National plant of the American

Smelting and Refining Co. 299
New Jersey Zinc Co 246

326

INDEX

Nutting, Mr.

PAGE

256

Ore and Fuel Co 63

different behavior of coarse

and fine in lead smelting 79
-treatment in detail by the
Huntington-Heberlein pro-

. . 150-155

Parkes process, cost of refining

by 99

Percy, Dr 244

Perth Amboy plant of the Amer-
ican Smelting and Refin-
ing Co 296

Petraeus, C. V 24

Pfort curtain for furnaces 82

Picher Lead Co 197

Piddington, F. L 263

Potter, Prof. W. B 15

Pueblo lead smelter 294

Smelting and Refining Co. 84
Pufahl, O. 38, 291, 294, 296, 299, 302,

304
Pyritic smelting without fuel

practically impossible . . . 256

Raht, August 251, 254

Refining, monthly cost of per

ton of bullion treated. . . 100

Roasters, hand, and mechanical
furnaces, average monthly
cost of 98

Roberts-Austen, W. C 139

Salts, effect of crystallization of

contained on concrete. . . 243
Santa Fe Gold and Copper Min-
ing Co 255

Savelsberg, Adolf 122

Savelsberg process 186-192

process, claims of in patent 201
process contrasted with Hun-
tington-Heberlein 209

process, difference between
and Huntington-Heber-
lein. . . 197

PAGE

Savelsberg process the simplest. 132
Scotch-hearth method, perma-
nency of 195

Scotch hearths 34

Schneider, A. F 81

Seattle Smelting and Refining

Works 273

Silver-lead blast furnaces, me-
chanical feeding of 81

blast furnace, proper condi-
tions 73

smelting, details of practice 73

smelting, modern 73

Slag-smelting costs 34

Slime analysis at Broken Hill. . 51
Slimes, analyses of in lead refin-
ing 281

desulphurization of by heap

roasting 51

treatment of at Broken Hill 53-

55
Smelter, new, at El Paso, Texas 285

Smelters' pay 32

Smelting, average cost of per ton 98

Smelting Co. of Australia 263

costs 48

detailed costs of 101, 102

of galena ore 38

preparation of fine material

for 59

Solution, washing from slime. . 277

Sticht, Mr 256

St. Joseph Lead Co 16

St. Louis Smelting and Refining

Co 81

Sulphide Corporation . . . - 145

Sulphur dioxide, effect of on

cement 240

Sulphuric acid, making of at

Broken Hill 174

Tasmanian Smelting Co 145

Tennessee Copper Co 255

Terhune, R. H., furnace gratings 84
Thacher, Arthur 14

Ulke, Titus.. . 270

INDEX

327

PAGE

United Smelting and Refining

Co 88

States Zinc Co 295

Vezin, H. A 252

Walls, retaining 237

Walter, E. W 260

Waring, W. Geo 24

Welch, Max J 229

Wetherill, Samuel 244

Wheeler, H. A.

PAGE
10

Zinc, amount required in lead

refining 265,266

crusts, treatment of in lead
refining 267

oxide in slags 108

retort residues, analysis of
materials smelted and
bullion produced 106

retort residues, smelting. . . 104

07564

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