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INDUSTRIAL CHEMISTRY

BEING A SERIES OF VOLUMES GIVING A
COMPREHENSIVE SURVEY OF

THE CHEMICAL INDUSTRIES

Edited by SAMUEL RIDEAL, D.Sc. Lond., F.I.C.

FELLOW OP UNIVERSITY COLLEGE, LONDON

ASSISTED BY

JAMES A. AUDLEY, B.Sc. J. R* PARTINGTON, D.Sc. (Vict.)

W. BACON, B.Sc., F.I.C. ARTHUR E. PRATT, B.Sc.

M. BARROWCLIFF, F.I.C. ERIC K. RIDEAL, Ph.D., M.A., F.I.C.

H. GARNER BENNETT, M.Sc W. H. SIMMONS, B.Sc.

F. H. CARR, F.I.C. R. W. SINDALL, F.C.S.

S. HOARE COLLINS, M.Sc., F.I.C. SAMUEL SMILES, D.Sc.

H. H. GRAY, B.Sc. D. A. SUTHERLAND, F.C.S.

H. C. GREENWOOD, D.Sc. HUGH S. TAYLOR, D.Sc.

•C. M. WHITTAKER, B.Sc.

INDUSTRIAL
ELECTROMETALLURGY

INCLUDING

ELECTROLYTIC AND ELECTROTHERMAL

PROCESSES

t BY

ERIC K^RIDEAL, MA. (Cantab.), Ph.D., F.LC,

JOINT AUTHOR WITH S. R1DEAL OP " WATER SUPPLIES

NEW YORK
D. VAN NOSTRAND COMPANY

25 PARK PLACE
1919

PRINTED IN GREAT BRITAIN

GENERAL PREFACE

The rapid development of Applied Chemistry in recent years
has brought about a revolution in all branches of technology.
This growth has been accelerated during the war, and the
British Empire has now an opportunity of increasing its
industrial output by the application of this knowledge to the
raw materials available in the different parts of the world.
The subject in this series of handbooks will be treated from
the chemical rather than the engineering standpoint. The
industrial aspect will also be more prominent than that of
the laboratory. Each volume will be complete in itself, and
will give a general survey of the industry, showing how
chemical principles have been applied and have affected
manufacture. The influence of new inventions on the
development of the industry will be shown, as also the
effect of industrial requirements in stimulating invention.
Historical notes will be a feature in dealing with the
different branches of the subject, but they will be kept
within moderate limits. Present tendencies and possible
future developments will have attention, and some space
will be devoted to a comparison of industrial methods and
progress in the chief producing countries. There will be a
general bibliography, and also a select bibliography to follow
each section. Statistical information will only be introduced
in so far as it serves to illustrate the line of argument.

Each book will be divided into sections instead of
chapters, and the sections will deal with separate branches
of the subject in the manner of a special article or mono-
graph. An attempt will, in fact, be made to get away from

vi GENERAL PREFACE

the orthodox textbook manner, not only to make the treat-
ment original, but also to appeal to the very large class of
readers already possessing good textbooks, of which there
are quite sufficient. The books should also be found useful
by men of affairs having no special technical knowledge, but
who may require from time to time to refer to technical
matters in a book of moderate compass, with references to
the large standard works for fuller details on special points
if required.

To the advanced student the books should be especially
valuable. His mind is often crammed with the hard facts
and details of his subject which crowd out the power of
realizing the industry as a whole. These books are intended
to remedy such a state of affairs. While recapitulating the
essential basic facts, they will aim at presenting the reality
of the living industry. It has long been a drawback of our
technical education that the college graduate, on commencing
his industrial career, is positively handicapped by his
academic knowledge "because of his lack of information on
current industrial conditions. A book giving a compre-
hensive survey of the industry can be of very material
assistance to the student as an adjunct to his ordinary text-
books, and this is one of the chief objects of the present
series. Those actually engaged in the industry who have
specialized in rather narrow limits will probably find these
books more readable than the larger textbooks when they
wish to refresh their memories in regard to branches of the
subject with which they are not immediately concerned.

The volume will also serve as a guide to the standard
literature of the subject, and prove of value to the con-
sultant, so that, having obtained a comprehensive view of
the whole industry, he can go at once to the proper
authorities for more elaborate information on special points,
and thus save a couple of days spent in hunting through the
libraries of scientific societies.

As far as this country is concerned, it is believed that
the general scheme of this series of handbooks is unique,
and it is confidently hoped that it will supply mental

GENERAL PREFACE vii

munitions for the coming industrial war. I have been
fortunate in securing writers for the different volumes who
are specially connected with the several departments of
Industrial Chemistry, and trust that the whole series will
contribute to the further development of applied chemistry
throughout the Empire.

SAMUEL RIDEAI,.

AUTHOR'S PREFACE

Amongst the many branches of applied chemistry electro-
metallurgy has shown a great technical development, and
in the following pages an endeavour has been made to
indicate both the limits and possibilities of the application
of electrolytic and electrothermal methods in this domain.

It is a matter of past regret and present concern that
Electrochemistry has not yet, except in one or two ex-
ceptional cases, been raised to the dignified standing of a
"subject" in our English higher educational system. As
a result electrochemical considerations receive but scant
attention, usually being confined to a few lectures in a
course covering the whole of physical chemistry.

Notwithstanding the fact that there exists an excellent
bibliography of text-books on theoretical electrochemistry,
this lack of a personal element in the teaching has been
sufficient to eliminate the English research student from
the field.

As a result the English Industry has suffered in having
either to purchase foreign processes or to waste both time
and money in experimental work carried out by investi-
gators ill equipped with the requisite knowledge. It must
not be forgotten that in the English-speaking countries
there are sources of power and raw materials in very great
variety, awaiting development at the hands of those who
are capable of taking part in the great post-war period of
reconstruction.

In many specific cases electrochemical processes offer
both economic, aesthetic, and industrial advantages over

x AUTHOR'S PREFACE

older chemical or metallurgical treatment, and in the de-
velopment of old processes or in the exploitation of new
ones, it is the hope of the author that the case for electro-
chemistry may have both due and deliberated consideration,
and that the present volume may contribute to this end.

In the sections on Electrolytic processes, the normal
hydrogen electrode has been taken as the arbitrary
standard of zero potential difference, whilst the conven-
tional positive sign is placed before the electrolytic
potentials of those metals which possess an electrolytic
solution pressure greater than that of hydrogen, i.e. those
elements which are most chemically active, whilst the
more noble elements are given a negative value on account
of their small solution pressures.

In many of the calculations the heats of formation of
certain compounds form the basis for the derivation of their
decomposition potentials. This method, as shown in the
introduction, although not strictly correct, usually gives
results sufficiently accurate for practical purposes where
direct experimental observation must necessarily include
small irreversible electrode effects. Special emphasis has
been laid upon the influence of colloids in electrolytic
deposition, whilst in electrothermal processes the dissocia-
tion of many stable compounds at high temperatures, and
the application of the partition' coefficient of substances
between immiscible slags and metals, have been used to
assist in the elucidation of the reactions involved.

E. K. Iv.

London,
June, 19 1 S.

CONTENTS

PAGE

GENERAL PREFACE v

AUTHOR'S PREFACE ix

INTRODUCTION.

The Ionic theory. Electrolytic potentials. Over-potential and passivity.
Reaction velocity. Diaphragms and cataphoresis. Fused electrolytes.
Power. The metalliferous resources of the British Commonwealth • . i

SECTION I.— ELECTROLYSIS IN AQUEOUS

SOLUTIONS.

Copper — Winning, refining, plating. Conditions for uniform deposition
of metals. Rotating electrodes. Complex electrolytes. Colloid
addition agents. Bronze and brass. Zinc. Galvanizing. Cadmium.
Gold. Parting of gold and silver. Silver. Lead. Antimony.
Bismuth. Tin. Detinning. Plating. Nickel. Cobalt. Cobalt-
nickel alloys 28

SECTION II.— ELECTROLYSIS IN FUSED

ELECTROLYTES.

Sodium. Potassium. Magnesium. Calcium. Strontium. Barium.

Lead. Zinc. Aluminium, aluminium alloys ..... 109

SECTION III.- THE ELECTROLYTIC PREPARA-
TION OF THE RARER METALS.

Gallium. Indium. Thallium. Cerium. Neodymium. Praseodymium.

Lanthanum. Boron. Vanadium. Titanium. Manganese. Uranium 135

xii CONTENTS

SECTION IV— ELECTROTHERMAL PROCESSES.

PAGB

Zinc. Copper. Nickel. Manganese. Chromium. Molybdenum. Tungsten.
Vanadium. Tellurium. Uranium. Zirconium. Silicon. Graphite.
Phosphorus. Arsenic. Carbon disulphide 139

SECTION V.— CARBORUNDUM AND OXYSILI-

CIDES OF CARBON.

Carborundum. Silundum. Silfrax. Monax. Siloxicon. Fibrox . .164

SECTION VI.— THE CARBIDES.

Properties of the carbides. Calcium carbide. Heat of formation. Block

or ingot furnaces. Tapping furnaces. Continuous furnaces . .172

SECTION VII.— ELECTROTHERMAL NITROGEN
FIXATION BY METALS AND METALLIC
COMPOUNDS.

The nitrogen problem. Arc processes, Hausser's method, the Haber pro-
cess. Biochemical methods The nitrides, cyanides and cyanamides • 183

SECTION VIII.— IRON AND THE FERRO-ALLOYS.

Electrolytic iron. Electrothermal pig iron. The production and refining
of steel. The functions of the slag. Dephosphorization, deoxidation
and desulphurization. Arc furnaces. Induction furnaces. Composite
furnaces. The "pinch effect " furnace. The high frequency induction
furnace. Ferro-alloys. Ferro-silicon, tungsten, manganese, chromium,
molybdenum, vanadium, titanium, uranium, and boron . . . 205

APPENDIX 237

INDEX . 239

INDUSTRIAL
~ ELECTROMETALLURGY

INTRODUCTION

The foundations of the general principles of electro-metal-
lurgy were laid by Michael Faraday in 1833, w h° introduced
the present nomenclature, e.g. such terms as electrolyte
electrode, cathode, anion, and gave us the fundamental
quantitative laws on which both the science and industry of
electrolytic processes are founded. It is frequently forgotten
that it is to Faraday we owe the important generalization
that a definite and unalterable quantity of electricity is asso-
ciated with each valency of an element, the first tangible
suggestion of the atomistic or electronic theory of electricity.

In 1853 Hittorf noticed that the concentration of the
solute in the solvent altered during electrolysis. The
concentration of copper sulphate, for example, in aqueous
solution increases at the anode and decreases at the cathode
when such a solution is subjected to electrolysis.

By a series of experiments he was able to calculate the
transport number of the ions, whilst Kohhrausch a few years
later (1869), by an elaborate investigation on the molecular
conductivities of dilute solutions, was able to determine the
values of the velocities of the ions under a definite potential
gradient. FromHittorf 's andKohlrausch'sfigures it is possible
to calculate the actual ionic velocities in cms. per hour under
a potential gradient of one volt per cm. in dilute solutions.
The following figures were obtained for solutions at 18 C.

Cations.

Anions.

Hio-8

OH' 5-6

K 2 05

cr 2-i2

Na i'i6

NO's i'9i

Ag i*66

2 INDUSTRIAL ELECTROMETALLURGY

The calculated figures were confirmed by actual measure-
ment of the migration velocity of the ions in solution by
Ix>dge, Whetham, Steele, and others.

Up to this stage in the development in the principles of
electro-chemistry no hypothesis as to the state of the solute
in the solvent was necessary. In 1887 the Grotthusian
hypothesis of electrolytic conduction was replaced by the
dissociation theory of Arrhenius and Van 't Hoflf. Accord-
ing to this theory a salt when dissolved in an ionizing solvent
is partially dissociated into free ions, according to the follow-
ing scheme —

Equilibrium is established in accordance with the laws of
mass action, and the solution as a whole is electrically
neutral.

A further advance was made in the subject by Van 't
Hoff, who applied the gas laws to substances in solution.
The concentration of the reacting constituents, undissoci-
ated molecules or ions in the solvent were regarded as
equivalent to the concentrations or partial pressures of gases
in a gaseous mixture.

There is no doubt that this conception has been of great

service to the electro-chemist, and the results obtained by

a rigorous application of the gas laws to dilute solutions

have been not only extremely varied, but of far-reaching

importance.

It is somewhat unfortunate that the development of

the ionic theory originally suggested by Arrhenius and Van
't Hoff was chiefly accomplished in Germany, where fre-
quently a somewhat pedantic train of thought tends to
• exclude other important factors from due consideration. In
this case the function of the solvent was entirely neglected
and treated rather as the convenient vacuum in which gases
could be distributed. Many discrepancies were noticed
between the experimental and calculated results, especially
when dealing with stroog electrolytes, and as a result in-
genious theories and formulae were proposed to square the

INTRODUCTION 3

facts either frankly empirical or based on some pseudo-
scientific generalization from Van der Waal's equation.

The experimental work of Walden (1904, et scq.) on the
conductivities of electrolytes in various solvents indicated
that the theory could not be retained in the simple form as
originally stated by Arrhenius and Van 't Hoff. At the
present time no new theory has been proposed capable of
the simple thermodynamic treatment which was one of the
great advantages of the old one, but at any rate we must
now consider the problem in a new light, as one in which
the solvent itself performs important, if not the most im-
portant, functions.

Ionization must be regarded as taking place subsequent
to solution (usually hydration) of the solute according to
the following scheme —

MX->MX(H 2 0) m ^M(H 2 0) rc +X , (H 2 0)

both the undissociated salt and the ions being surrounded
by envelopes of the solvent. The nature of the forces
holding the envelope round the solute, as well as the number
of solvent molecules in each envelope, is as yet a matter of
uncertainty, but it seems probable, from a consideration of
the ionic mobilities, that the ionic hydration numbers are
small, rising to 6 or 9 molecules per ion in the case of the
ions of small atomic weight, e.g. Uv or F', and may be entirely
absent in the heavier ones, such as Cs* and I'.

Electrolytic Potentials.

In addition to the ionic theory, the hypothesis of
electrode " solution pressure " advanced by Nernst in
1889 has been of great assistance in developing the
science. On this hypothesis, all metals possess a solution
pressure or a tendency to drive ions into solution. Since
the metallic ions leaving the metal are positively charged,
the electrons or negative charges are kept in the metal ;
metallic ions are consequently forced into solution until
the potential difference between the metal and layer of

4 INDUSTRIAL ELECTROMETALLURGY

solution in contact with the metal is great enough to prevent
the further discharge of metallic ions.

Imagine the transfer of 8n gm. ions of a v valent metal
to pass from the electrode to the solution. The electrical
work is evVSn where ve8n is the charge carried by 8n gm.
ions. This transfer is also equivalent to bringing hn gm.
ions from the solution pressure P to the solution of osmotic
pressure aC, where C is the concentration of the salt in the
solvent and a its degree of ionization, and equal to

8«RT log -^
aC

By the principle of virtual work,

p PT P

V^8w=8«RT log X or V= — log f-

aC ve aC

We can also arrive at a similar relationship in the
following manner : —

If a metal of valency v be placed in an electrolyte con-
taining its ions, and equilibrium is established when the
difference of potential between the metal and solution has
risen to a value V and the ionic concentration of the metal
in the solution has risen to aC, and further, if fa and [l be
the chemical potentials of the ions in the solution and the
uncharged molecules in the metal respectively, with an
electric charge ve on each ion, equilibrium is established

when

fa— fx—Vve

But fa^+RT log aC

hence Vas _ft-ri+RT logaG

ve
putting /*— /x =RT log K

VRT 1 K
ve aC

where K is to be regarded as the solubility constant of the
metal in the form of metallic ions.

The electrolytic solution pressures of the metals as
calculated from the measurements of the electrode potentials

INTRODUCTION 5

vary very considerably ; for example, P for zinc is equal to
io 17 atmospheres and for palladium equal to io- 31 atmo-
spheres.

It is evident that although the conception of electrolytic
solution pressure is a convenient one, it cannot be a true re-
presentation of the facts, and it would appear more reason-
able to adopt Smit's and VanLaar's suggestion of replacing
the term solution pressure P by K, the solubility constant
of the metal in the form of its metallic ions. In those cases
where the electrode is composed of an alloy or amalgam of
two or more metals, the theoretical calculation of the electro-
lytic potentials has been made by Van Laar,* to whom the
reader is referred.

In the following pages the electrolytic potentials of the
metals in a solution of normal ion concentration are all
referred to the normal hydrogen electrode, which is taken
at the arbitrary value

p __ RT - P hydrogen at i atmosphere __
h ~~ e C normal hydrogen ion solution "~~

The determination of the true value of electrolytic potentials
is a matter of some difficulty, since a zero E.M.F. between
a metal and solutions of its salts cannot readily be obtained ;
but the dropping electrode of Paschen and Palmaer f may
be considered as being the most satisfactory attempt
to devise an auxiliary electrode comprising electrode and
electrolyte of zero potential difference.

The calculation of the potential difference between two
electrodes in the same or different electrolytes separated
by a diaphragm, by means of the equations developed by
Nernst, Henderson and Planck, can be obtained not only
from a knowledge of the respective electrolytic solution
pressures of the metals, the concentrations of the solutions
and the mobility of the respective ions, but also as shown
by Helmholtz and Thomson in 1847, from a knowledge
of the heat of reaction.

If we consider the simple system, copper/copper sulphate/

* Elektrochemie, Amsterdam, 1907. f ZHt. Phys. Chem., 25, 265, 1895.

6 INDUSTRIAL ELECTROMETALLURGY

zinc sulphate/zinc, the zinc and copper being joined by a
wire, and imagine it at work at t * until 8n gms. of zinc are
dissolved and copper deposited, we then raise the tem-
perature to t+8t and pass a current through the cell so as
to redeposit the 8n gm. ions of zinc and dissolve the same
amount of copper, subsequently allowing the cell to cool
again to t

If the E.M.F. of the cell at t is n volts, at t+8t, tt-Stt,
the electrical work done by the cell is equal to ve8nn, where
v is the valency of the metal, in this case 2, and e is the
charge per gm. equivalent of a monovalent element ; whilst
the chemical work is equal to 8nq> where q is the heat of
solution of a gm. atom of zinc — the heat of solution of a gm.
atom of copper.

The energy given out by the cell is consequently 8n(q
•—eirv). The heat absorbed at the higher temperature will
be in a similar manner equal to8n{q—ve(ir+8ir)) if the heat
of reaction does not change sensibly with the temperature.

During the cycle, the external work performed is equal
to vSneSn and the quantity of heat given out at the lower
temperature is 8n(q— etro).

N v8ne8iT __8n(q—e7Tv)

ow — ^

q , .877
or 7T = -L + 1 —

ve 8t

8it
In many cases the temperature coefficient - is so small

8t
that the term t— may be neglected for approximate calcu-

lations of n.

OVERPOTENTIAI, AND PASSIVITY.

With a potential difference smaller than that calculated,
the passage of the current is only associated with concen-
tration changes in the electrolyte, provided of course that

* / is measured in degrees on the absolute temperature scale.

INTRODUCTION 7

other ions capable of being discharged at the lower potential
are not present in the electrolyte. For example, in the
electrolysis of dilute sulphuric acid between platinum
electrodes, the following ionic discharges take place with
increasing applied potential difference.

P.D. in volts. Ionic discharge.

ro8 {nrvn-i

»2

o"-»o.

167 aOH'->H f O+0,

' Ih-»h 2

S0 4 "-»H 2 S0 4 +0 2

1 "95
2*60

H-»H S

HSO/-»HS0 4 and H 2 S 2 0,

2-83 I H->H 2

1 30"->0 3

In practice a potential difference considerably higher
than that calculated has to be applied to bring about
electro-deposition at an economic rate. The uses of de-
polarizing agents added to the electrolytes in order to
reduce the applied voltages necessary for electrolysis, and
thus lower the electrical energy consumption at the expense
of the depolarizer, will be dealt with in subsequent sections.

Frequently the excess potential difference found neces-
sary can be traced to the occurrence of irreversible pheno-
mena taking place at the surface of the electrodes. Apart
from the general one of the Joule heat loss due to the re-
sistance of the circuit, those causes of discrepancy between
theory and practice may be accounted for by one of the
following factors : —

Overpotential. — In those processes where cathode
hydrogen is liberated, it has been noted that the P.D.
necessary for hydrogen liberation, when the same electrolyte
is used and identical anodic reactions take place, is not
independent of the nature of the cathode. The theoretical
applied E.M.F. has always to be increased by a certain
definite amount for each particular metal. According to

8 INDUSTRIAL ELECTROMETALLURGY

Caspari, the following are the values of the over-

potential i\ : —

Metal.

1} in volts.

i '3

i*3

1*22

115

I '00

o-8o

0*19

Pt (bright)

0*07

Pt (black)

O'OO

No satisfactory explanation for this phenomenon is as
yet forthcoming, although attempts have been made to
correlate the t\ values with the heat of formation of hypo-
thetical hydrides, with surface tensions, with the formation
of absorbed gas and with the diffusivities of gas molecules
and the gas ions in the metal. Advantage is taken of the
high overpotentials exhibited by certain metals in certain
electrolytic reduction processes and in the technical deposi-
tion of zinc and cadmium.

Passivity. — As in the case of cathodic hydrogen, the
anodic evolution of oxygen is also associated with irrevers-
ible overpotential phenomena, usually, however, of quite
inconsiderable magnitude A more serious disturbance of
anodic processes is the occurrence of passivity. In certain
electrolytes metals may exhibit no tendency to anodic
solution ; the metal appears more " noble " than is actually
the case. Electrolytes containing oxidizing acids are more
prone to cause this phenomenon than others, and although
probably all metals may be passified by suitable treatment,
the following exhibit the characteristics to a marked degree :
iron, aluminium, cobalt, chromium, platinum, tungsten, and
molybdenum. Various theories have been proposed to
account for the phenomenon of passivity, which have been
summarized in the "Transactions of the Faraday Society "
for 1916 ; amongst the more important may be mentioned :

I . The formation of a gas film on the surface.

INTRODUCTION 9

2. The formation of an oxide film. ,

3. The conversion of the surface metal into an allotropic
modification, of which the electrolytic solution pressure is low.

4. The velocity of ionization or hydration of the ion is
retarded below its normal speed.

Reaction Velocity. — The maximum speed at which
any change involving a cycle of operations may be made tfo
take place is set by the maximum velocity of the slowest
intermediary link. This generalization is frequently over-
looked in electro-chemical process, but is nevertheless one
of the most important factors in the cause of electrical
inefficiency, as may be indicated by the following examples :

If we use an alternating current to perform the electro-
lysis of copper sulphate with copper electrodes, we can
imagine that when one electrode becomes the anode the S0 4 "
ion will be discharged, forming cupric sulphate with solution
of the metal ; at the next instant the current is reversed, and
the cupric ion will be discharged The sum total of the two
reactions can be represented as follows :

Cu^Cu"'

It is evident that no change in weight of the electrode
should take place if the ions simply oscillate to and fro from
the electrode to the solution, but if they are removed whilst
in the solution by hydration, unless hydration occurs
instantaneously, a net loss in weight will result. Leblanc
found by experiment that the alternations of a current of a
periodicity of 50 and above were sufficiently quick to
prevent such a loss, whilst the ions were in solution, but
for less frequent alternations solution did actually take
place. If potassium cyanide were present in the electrolyte,
solution occurred up to 500 periods, but ceased at 10,000.
From these figures it is clear that the rates of transforma-
tion of the various modifications into each other, which
metallic copper has to uniergo before it passes from the
anode to the cathode in an electrolytic cell, are by no
means instantaneous, and do actually set a limit to the
velocity of electrolysis.

io INDUSTRIAL ELECTROMETALLURGY

* Electrolysis in Fused Electrolytes and Electro-

thermics.

The isolation of the alkali metals by electrolysis of the
fused hydroxides, and the technical method of manufacture
adopted at the present day, was first accomplished by Sir
H. Davy (1800 to 1810). A systematic investigation of the
properties of the fused salts has been made by Lorenz and
his pupils, who showed the general applicability of Faraday's
laws to these electrolytes. In general the conductivity of
fused solutions is much superior to aqueous solutions, but
at the same time, owing to various disturbing influences,
the current efficiency is usually lower. The following are
the more important causes of low efficiencies : —

1. Evaporation of the deposited metal. — A very con-
siderable loss may occur due to vaporization of the de-
posited metal. This factor becomes increasingly important
the higher the melting point of the metal, since the
temperature interval through which the liquid exerts an
appreciable vapour pressure is always larger for metals of
low melting point.

2. Chemical side reactions. — In fused electrolytes the
intermediary formation of sub-salts unstable in aqueous
solutions is of somewhat frequent occurrence. A notable
example is found during the electrolysis of calcium chloride,
where the cathode formation of the coloured subchloride,
CaCl, causes a reduction in the yield of metal.

3. Cloud formation. — During electrolysis of certain
metals, especially lead, in fused electrolytes containing
alkalis, the precipitated metal will frequently not coalesce,
but is dissipated through the electrolyte in the form of a
fine cloud or mist. It is as yet uncertain whether the cloud
consists entirely of the metal cathodically deposited in the
form of a colloidal solution or whether it contains a small
quantity of the alkali metal, alloyed or combined with it.
To obviate or minimize cloud formation, the temperature
of the electrolyte should be maintained as low as possible.

4. Solution of the Metal in the Electrolyte. — As in

INTRODUCTION n

the case of aqueous electrolytes, the determination of the
electrolytic potentials of the metals in fused solutions
has been attempted, but no high degree of accuracy was
obtained owing to experimental difficulties in connection
with the construction of a suitable auxiliary cathode, and
the rapidity of diffusion of the fused electrolytes. The
values of the electrolytic potentials are usually approximated
by interpolation from the figures obtained from amalgams
of known composition in aqueous electrolytes.

The development of electro-thermal processes has, as is
indicated by its name, been confined to the chemical effects
produced by the Joule heat liberated by the passage of the
current through resistances either of the first or second
class. The upper temperature limit obtainable in an electric
furnace is that temperature at which the rate of sublimation
of carbon becomes appreciable and has been estimated at
from 3000 C. to 3600 C. By this means the preparation
of a number of high-temperature products hitherto unpro-
curable has been a matter of no great difficulty, whilst the
efficiency of high-temperature smelting has increased hand
in hand with the simplification of the operation.

Diaphragms and Cataphoresis.

During the last few years the electrical properties of
colloids have received ever-increasing attention by investi-
gators in the subject of physical and electro-chemistry,
which cannot fail to be reflected in the electro-chemical
industry of the future. A discussion of the results already
obtained lies somewhat outside the province of this volume,*
but it may be remarked in passing that at the present
time there are three distinct lines of research in this field
which have already proved extremely helpful in the
industry : —

1. The preparation of colloidal metals as sols in various

* For further information on these subjects, the reader is referred to :
Svedberg, " Die Methoden zur Hersteliung Kolloider Losungen." V. Wei-
marn, " Grundzuge der Dispersoid Chemie." Freundlich " Kapillarchemie."
Donnan, Membrane equilibria, Zeit.f. Elektrochemie, 17, 572, 191 1.

12 INDUSTRIAL ELECTROMETALLURGY

dispersion media by the two methods, (a) cathodic dispersion,
and (b) dispersion by means of an electric arc.

2. The use of protective colloids in the electrolytic de-
position of metals.

3. The calculation of the drop of potential across dia-
phragms, and also the velocities of ionic migration through
the pores of the materials used in electrolytic operations,
where it is desirable to separate the anode and cathode
compartments.

Power.

The deciding factors in the choice of a suitable method
for the preparation of any product on a manufacturing scale
are generally exceedingly complicated, and the relative
values of raw materials, energy, labour and transportation
to market costs vary from country to country, and, indeed,
from place to place. The ideal site for a factory in a given
locality cannot be indicated on a map by a strictly scientific
method, such as marking off the power source at a point A,
raw material sources at B and C, the distribution centre at D,
and calculating the position of the site. Practical experience
has shown that the energy factor is all important, and
that big industrial enterprises spring up round the power
sources.

Industrial electro-chemistry requires its power in elec-
trical form, and the values of energy in this form are the
dominant factors for the formation of these industries.
The two chief sources of power are water and coal, although
solar radiant energy, gasified peat and turf, and various
organic fermentation processes, are being utilized on an in-
creasingly extensive scale. It has been frequently claimed
that owing to the relative cheapness of water - produced
electricity compared to coal, no electro-chemical industries
in a coal-producing country can compete with countries rich
in water power. Many figures given for the actual cost of a
kilowatt year are fallacious, owing to the different conditions
obtaining both in supply and consumption of the energy
produced and delivered from a generating station.

INTRODUCTION 13

If the current be used solely for heating, as is the case
in many electro-thermal operations, then owing to the
heavy outlay necessary to build a water-power generating
plant, the electrical energy derived from water is roughly
about ten times as expensive at the present time as the
fuel energy in coal. If, however, electrical energy is
required in both cases, the expense of converting fuel into
electrical energy usually makes it the more expensive of
the two.

The running expenses of the two plants are also widely
different ; the fixed cost in a fuel-driven generator may be
taken at about half the total cost of production, whilst in a
water-driven plant the fixed cost is practically the only one
to be considered. In order to arrive at comparative figures,
it must be remembered that the efficiency of water-driven
plants has practically reached its upper limits, whilst the
inflation by prospectors and real estate agents of the cost of
possible water-power sites as well as the growing aesthetic
public opinion against the destruction of nature's scenery,
are all tending to elevate the water-produced power costs.
On the other hand, the improvements in turbine-driven
plant, gas firing, utilization of coal by-products, and the
possible advent of the gas turbine may lower the cost of the
fuel-produced electricity. At the present time electricity
produced in an up-to-date generating station can compete
quite favourably with water power in those areas which
have developed large industries like Niagara, but cannot
compete with the newer electrical countries, such as Norway,
Alaska or Africa, where land and water costs are small.
In these cases the cost of importing the raw materials and
the export of the manufactured article bid fair to com-
pensate for the cheaper power costs.

The general conditions necessary for bringing down the
power costs to the minimum are common to both methods
of production. Economy in the production of electricity
depends entirely upon the continuity of production and
consumption, i.e. the load factor. Many electricity genera-
tors try to sell their interpeak current at a low rate to

14 INDUSTRIAL ELECTROMETALLURGY

flatten their production curve, whilst the power consumers
in their turn desire cheap power but cannot afford to buy
discontinuous current, which is liable to be cut off at any
moment by the producers, except in special cases for electric
furnace work. At the present moment the problem of link-
ing up various power stations to avoid clashing of peak
currents in the areas using electrical power is being con-
sidered, but its practical operation is one of considerable
difficulty. Economy can only be effected by producers
supplying a large amount of power to consumers at a steady
rate, and the amount must be so great that the lighting and
heating load for the labour in the industries must be but a
small proportion of the total load. Under these conditions
of production and consumption the problem of power
transmission becomes important. The present tendency
leans to high voltage transmission, using copper or aluminium
lines.* The first attempts to transmit'at 10,000 volts were
made in 1891. In 1901 50,000 to 60,000 volts transmission
lines were in operation, and at the present time 110,000 to
150,000 volt lines are constructed. By raising the voltage
increased quantities of power can be transmitted for longer
distances at minimum cost for the conductor. There are,
however, limits both to the voltage employed and to the
distance over which the power has to be transmitted. On
raising the voltage, not only have we the same Joule (Watt
loss) loss by heating on the line, but the corona loss (above
100,000 volts) increases rapidly. The critical voltage at
which the corona loss commences depends on the tempera-
ture pressure and presence of dust or fog in the air, as well
as the radius of the line conductors. The smaller the
conductor, the earlier does this electric brush discharge
commence. Again, with high voltage lines, the expense of
substituting steel or ferro-concrete towers for wood pole
lines, and the fitting of high voltage insulators, raises the cost
of line construction.

* Although the specific conductivity of aluminium is only one half that
of copper, yet per unit of weight, aluminium is a better conductor. Metallic
sodium protected by glass tubes has also been suggested as a possible
conductor.

INTRODUCTION 15

These factors, which are specially important in industrial
areas where precautions have to be taken, must be considered
when the question of installing a separate power plant or
the transmission of power from some distant generating
station is considered. In western America the limiting
distance appears to be in the neighbourhood of 250 miles,
and in the east about 150 miles. If the station is further
away it then becomes more economical to instal one's own
generating plant. In England, where the population is
denser and the cost of transmission lines considerably
enhanced, the economical distance of transmission would
be still shorter.

The Interim Report on the Electric Power Supply in
Great Britain, April 1917, points out the necessity for the
development of very large power centres ; the average of
some 600 undertakings in Great Britain have power stations
of 5000 h.p., or about one-fourth the capacity of one single
generating machine of economical size, and about one-
thirtieth of the size of what may be considered as an eco-
nomical " power-station unit." Thirteen of such " super-
power " stations are contemplated.

The present methods of power production can be
divided into the following groups : —

1. Hydro-electric.

2. Coal.

3. Gas.

4. Oil.

Hydro-electric Power. — Water power is exceedingly
scarce in Great Britain, the only large installation being at
Kinlochleven, where about 20,000 kw. is developed for the
production of aluminium ; potential sites may be found
both in Wales and Scotland. Within the Commonwealth
there are several very large undeveloped power sources,
notably in Ireland on the Erne and Shannon, in Canada,
Egypt, India, South Africa, New Zealand and British
Guiana. The cost of installation and running vary very
considerably from country to country. The following

16 INDUSTRIAL ELECTROMETALLURGY

figures may be taken as the approximate pre-war running
costs : —

Place. Total cost per kw. year to consumer.

Kinlochleven . . . . . . 45s. 8d.

Ontario, Canada
Hora-Hora, New Zealand
N. California, U.S.A.
Saulte-Ste. Marie, U.S.A.
Niagara, U.S.A. side
I^egnano, Italy
Turin, Italy
Brian9on, France

41s. 3^.

81s. yd.

68s. $i.

54s. 6d.

67s. id. to 112s. 8d.

20s.

545. 4d.

18s. 2\d.

The Scandinavian development of water power during
the last few years has been a remarkably large one. Very
low figures are quoted for the cost at Odda and Svaelgfos
in Norway and Trolhatten in Sweden, e.g. from ns. to 12s.
per kw. year. The actual selling costs of this power, when
full running costs, depreciation, and Government royalties
are included, are higher. The Norwegian and Swedish
Governments seem to be of the opinion that the natural
economic selling costs lie between 25s. and 40s. per kw.
year.*

Although the figures cited above are subject to wide
variations, the approximate figure of 40s. per kw. year may
be taken as a fair average selling price, on a pre-war basis,
for hydro-electric power, where the installation costs are not
high.

As is naturally to be expected, the pre-war installation
costs for hydro-electric power vary widely, depending on
the size of the plant and the engineering difficulties asso-
ciated with the erection. In the U.S.A. £26 seem to be taken
as a conservative standard cost, whilst the Kinlochleven
installation in Scotland is said to have cost £27 per kw. In
Norway the installation costs average about £15 per kw.,
rising to over £20 per kw. in the later installations. The
earlier plants were installed at a much cheaper rate, owing
to the fact that a large choice of available sites was per-
missible.

* Norwegian Royal Commission, Sept. 191 5.

INTRODUCTION 17

The feasibility of using tidal energy has been discussed
from time to time. This potential source of energy suffers
from the disadvantage that to ensure continuity of supply
large reservoirs would have to be erected to deal with the
periods of slack water.

Coal Power. — A pound of good quality coal will pro-
duce from 11,500 to 14,000 B.T.U., and the average may be
taken as 13,000 B.T.U.s per pound ; since a kw. hour is
equivalent to 3415 B.T.U.s, an ideal engine should be able
to produce 4 kw. hours of electrical energy per lb. of coal.
The most efficient steam-driven generator at present existent
is undoubtedly the turbo generator, which offers the ad-
ditional advantages of having low maintenance changes,
and can be constructed in large units at very cheap in-
stallation costs. For large turbo generators, units of 20,000
to 50,000 kw. capacity, which are considered to be the
minimum sizes compatible with economic working effici-
ency, about 15,000 B.T.U.s would be required per kw.
produced. Smaller plants at present in operation con-
sume some 20,000 B.T.U.s per kw. hour. The thermo-
dynamic efficiency of such a generating set would therefore
be 227 per cent. The capital installation costs for the
large units are estimated to lie between £11 and £13 per
kw. installed, figures which compare extremely favourably
with those cited for the hydro-electric power installa-
tions.

In 1915, 253,179,000 tons, and in 1916, 256,348,381 tons
of coal were mined in Great Britain, of which about one-
quarter left the country. It is evident that a slight export
tax on such a valuable raw material would considerably
lower the price for the home consumer. It is stated that *
the average selling price of coal in 1914 was 9s. irygd., and
in 1915, 12s. 5'6orf. per ton. It would,, therefore, appear
possible to deliver coal in bulk at the large power stations
contemplated by the reconstruction committee at from
7s. 6d. to 10s. per ton. Taking ys. 6d. as the m inimum we

* " Mineral Production of the United Kingdom in 1915 : Mines and
Quarries." Pt. iii., c. 8444.

I,. 2

18 INDUSTRIAL ELECTROMETALLURGY

obtain the following minimum cost of production per kw.
year : —

Coal .. .. .. .. .. .. 33s. zod.

Running costs and depreciation =10 per
cent, on Installation cost . . . . 21s.

Total . . 54s. xod.

To produce 1 kw. year for 40s., the pre-war value of
hydro-electric power, coal would have to be delivered at
the power plant for 4s. 3d. per ton. Although the post-war
price for hydro-electric power may be considerably higher
than the pre-war rate, yet a still greater increase in the cost
of raising steam must be expected. It is extremely probable,
however, that the altogether disproportionate rise in the
freight rates will be sufficient to swing the pendulum over
to the side of those power installations which are situate
close to their markets.

Gas Power, — The boiler efficiency of a plant where
steam is raised by gas firing is some 5 per cent, better than
where coal is employed, owing to the fact that in the one case
the fuel is perfectly homogeneous, requiring a definite and
fixed quantity of air for combustion, whilst in the case of
coal firing, combustion proceeds in stages, necessitating a
variable air supply ; thus, liberation of unburnt fuel as
smoke, with the deposition of partly carbonized tar on parts
of the heating system, can scarcely be avoided. Apart
from this consideration, the market value of the by-products
obtained in the distillation of coal is greatly above their
value as fuel suitable for raising steam, and it would appear
economically sounder to recover these by-products even if
their fuel value were lost to^the power plant. This would
naturally necessitate an increased coal consumption as far
as electrical pow^r production was concerned, but on the
other hand countries with supplies of coal available would
obtain large quantities of products useful as raw materials
for their various industries.

The relative advantages and disadvantages of partial
gasification of the coal at low or high temperatures, under a

INTRODUCTION 19

slight pressure or vacuum, lie outside the province of this
book. It is evident, however, that the nature and amounts
of the various by-products being dependent on the conditions
of gasification, can be controlled so as to suit the conditions
of the market for fertilizers, benzol, chemicals, metallurgical
coke, power or illuminating gas. We will only consider the
hypothetical case where the coal is practically completely gasi-
fied to produce gas, by-products and clinker in one operation
by the suitable introduction of steam, which we will assume
can be obtained as waste from the main generating plant.

The efficiency of a gas producer is in the neighbourhood
of 70 per cent. One ton of coal, containing 30 million
B.T.U.s., on complete gasification will give nearly 60,000
cubic feet of gas containing 21 million B.T.U.s.

We have further noted that 15,000 B.T.U.s are required
for a coal-fired boiler to give 1 kw. hour ; with a 5 per cent,
better efficiency for gas firing, 14,250 B.T.U.s would be
required. Hence for a kw. year, 5*96 tons of coal would
be required With coal at 7s. 6d. per ton, a kw. year with
a gas-fired turbo-generator would cost : —

For coal . . . . . . . . . . 44s. 8d.

For gasification 32s. od.

Running costs and depreciation . . 21s. od.

Total . . 97s. 8d.

or 42s. xod. dearer than a coal-fired turbo-generator set.

Against this must be set the value of by-products obtained
in gasification of coal, amongst which may be mentioned : —

Present net * value.

Rectified tar . . . . 41s. yd. (If sold as crude tar

8s. zod. 20s. g$d.)

Ammonium sulphate
Benzol
Sulphocyanide
Sulphuric acid
Pan coke and breeze

Total

5s. nji.
is. 3jd.
2S. 4%d.
3s. 10S.

63s. io\d.

* Net value represents possible profit on sale at current rates, after
d educting working costs for recovery and making allowance for depreciation
of any special machinery utilized.

20 INDUSTRIAL ELECTROMETALLURGY

Under these somewhat idealistic conditions the price
of a kw. year would fall from 97s. 8d. by 63s. lod. =335. iod.,
or 65. 2d. below the average water-power costs. If large,
installations were erected for the gasification of coal in
connection with the turbo-generator stations contemplated,
doubtless the price of by-products on the market would
fall and the disparity between costs of the alternate methods
of production would tend to disappear, but there can be no
doubt as to which system would be most advantageous to
the nation.

Gas Engines. — We have already noticed that to pro-
duce 1 kw. hour in a coal-fired turbo-generator, 15,000
B.T.U.s are required, which figure may be reduced to 14,250
B.T.U.s if gas firing is adopted. In a good modern large
gas engine only 13,500 B.T.U.s would be necessary to
produce 1 kw. hour. In addition, by the use of the heat
from the exhaust, steam can be raised to run a subsidiary
plant. The exhaust heat of 13,500 B.T.U.s of gas exploded
in the gas engine will raise steam equivalent to 3,500 B.T.U.s.
In other words, the net energy consumption per kw. hour is
only 10,000 B.T.U.s. In spite of the apparent advantages
in the use of a more efficient prime mover, the limitations
in size, 3000 kw. units being the largest constructed,
together with the heavy installation expenses, make the
working costs and depreciation on machinery more than
counterbalance any fuel economy which is to be obtained.

The natural solution for power generation by means of gas
is the realization of that long-sought machine, the gas turbine.

Power-gas sources are to be found in peat, turf, natural
oil, gas wells, and in various fermentation industries, such
as in the production of acetone, the retting of flax, and the
hydrolysis of sewage in septic tanks.

Oil Engines. — The Diesel oil engines are the most
efficient prime movers in technical operation, the B.T.U.
consumption per kw. hour being only some 8,500. For
relatively small electro-chemical installations, a Diesel
engine operating on a low grade oil or gas-works tax would
present several advantages.

INTRODUCTION 21

The Metaujferous Resources of the British

Commonwealth

In subsequent sections of this volume a short descrip-
tion is given of the various electrolytic and electro-thermal
methods employed for the isolation of the metals and the
production of metallic alloys and compounds. We have
already indicated that in the British Commonwealth, in-
cluding England, there exist the potential sources of large
quantities of electrical energy, capable of being produced
at low rates. For the successful development of a thriving
electro-chemical industry, a few conditions only need be
observed. Firstly, the necessary enterprise of financiers
and manufacturers ; secondly, the development of the
educational system so as to ensure the supply of specialists
trained in the branches of electro and physical chemistry,
without which knowledge no old process can be economically
modified, or new one developed, so as to compete in the open
market with the highly organized foreign undertakings.
The third important factor, viz. the availability of cheap
electric power, has already been discussed. We have
still to consider the possible lack of raw material for those
industries which we must create to ensure our economic
stability. The errors committed before the war, typified
by such glaring examples as allowing the control over the
Broken Hill Australian zinc ores and the Brazilian and
Travan£ore Monazite sand deposits to be taken over by
Germany, will probably not be repeated. On the other
hand, other nations, profiting by these examples, will become
more appreciative of the value of their own deposits, and the
supply of raw materials from foreign countries will be partly,
if not entirely, replaced by offers to supply manufactured
goods. We shall, therefore, be compelled to return to our
natural resources, if, indeed, our national spirit has not been
sufficiently aroused by recent events, so that we shall prefer
to develop our own resources even if foreign offers appear
more advantageous*

The location of deposits of metalliferous ores within the

22 INDUSTRIAL ELECTROMETALLURGY

British Commonwealth given below are drawn from the last
Report of the Advisory Council to the Department of Scien-
tific and Industrial Research,* and a paper by C. Cullis to
the Society of Engineers.! It will be noted that the supplies
at present available and capable of being developed at some
future date are by no means inconsiderable, and if stock of
the world's ore deposits could be taken, it is probable that
the greater portion of the metalliferous ores, with some few
exceptions, would be found to be located within this area.

Iron. — The quantity of iron ore smelted in the United
Kingdom in 1913 was 24 million tons, of which 8 million
tons were imported. Large deposits of iron ores are found
in the following countries : —

Great Britain. — Hematite, magnetite and ironstone in
many counties.

Scotland. — Ironstone in Ayr, Dumbarton, Fife, Lanark,
Linlithgow, Midlothian, Renfrew and Stirling.

Ireland. — Ferriferous bauxite in County Antrim. Hema-
tite in County Down, County Wicklow, Cork, Clare, Long-
ford and Leitrim.

Newfoundland. — Hematite on Bell Island.

Canada. — Hematite in Nova Scotia, Ontario and the
Yukon. Magnetite in New Brunswick, Quebec and British
Columbia.

India. — Hematite and ironstones in the Bengal Presi-
dency and the Central Provinces. Magnetite in the Madras
Presidency.

South Africa. — Siliceous hematite in W. Griqualand
and Bechuanaland. Siliceous magnetite in the regions
round Pretoria. Low-grade ore is stated to be plentiful in
Rhodesia.

Australia. — Hematite in South Australia, New South
Wales, Victoria, and parts of Western Australia and Queens-
land. Iron ore in Western Australia, a few miles north of
Perth.

Tasmania. — Magnetite, estimated at 25 million tons, is
stated to be available.

* Published May 191 7. f Trans. Sec. Eng., Dec. 1916, p. 25.

INTRODUCTION 23

New Zealand. — The Parapara deposits of limestone and
the magnetite deposits at New Plymouth are reported to be
extensive.

Chromium. — The chief exporters of chromite within the
British Commonwealth, according to the most recent returns,
were : —

Country. Date. Metric tons.

Rhodesia 1913 . . 63,384

Canada . . . . . . 1915 . . 11,486

India 1914 . . 5,888

Australia 1915 . . 638

Deposits also occur in Scotland, the Transvaal, New-
foundland and New Zealand.

Cobalt. — Up to 1904, New Caledonia supplied 90 per
cent, of the world's output. In that year the development
of the Ontario silver cobalt nickel mines began, and these
have now obtained the monopoly in the production of
cobalt. The Commonwealth producers of cobalt ore are : —

Country. Date. Metric tons,

Canada 1914 . . 401

N. S. Wales . . . . 1910 . . 10

Other sources of supply may be f ound in India at Jaipur
and in the Balmoral district of the Transvaal.

Manganese. — According to the Home Office statistics,
the production of manganese ores was : —

Country.

Date.

Metric tons.

United Kingdom

. . 1915

4,640

India

. . 1914

. . 693,824

Canada

. . I915

Queensland

• • 1915

203

Extensive deposits are also found in Egypt, New Zealand,
Newfoundland, Cape Colony and the Gold Coast.

Molybdenum. — The following was the world's produc-
tion of molybdenite in 1915 : —

Country. Metric tons.

N. S. Wales 35

Queensland . . 99

Canada . , . . . . . . . . 128

INDUSTRIAL ELECTROMETALLURGY

Molybdenum is also found in England, Scotland and
India.

Nickel. — The production of nickel ore is practically
confined to the extensive Cobalt and Sudbury deposits of
Ontario. In 1915 the amount of nickel ore mined in this
area was over 1,300,000 tons. Other deposits in Canada
are said ta exist, notably in Northern Alberta. Nickel has
also been reported to be present in deposits in East Griqua-
land, S. Africa.

Titanium. — Deposits of rutile and titaniferous iron ore
are found near Quebec, Canada ; New South Wales and near
Adelaide, S. Australia, as well as in New Zealand, but up
to the present time have not been developed.

Tungsten. — The production of tungsten ores within the
Commonwealth during 1914-1915 was as follows : —

Country.

Metric tons.

United Kingdom

Burma

Malay States
Canada

. . 329

.. 2,326

291

Queensland

N.S.Wales

.. 663
.. 83

Victoria (1913)

New Zealand (1913)

Tasmania (1913)

50
231

.. .. 58

Deposits which have not yet been worked are found at
Rajputana in India, and on the Subti river in Rhodesia.
Before the war over one-half of the world's tungsten ore
consumption was mined within the Commonwealth, yet no
ferro-tungsten and but a very small quantity of metallic
tungsten was manufactured in England. Ferro-tungsten
is now produced at Widnes, Luton and Sheffield, in England,
for the English Steel Industry.

Vanadium. — Small deposits of mottramite (Pb and Cu
Vanadate) have been observed in England in Cheshire,
Wiltshire and Shropshire, but appear to be too small to be
worked on a commercial scale. It is stated that Broken
Hill, Rhodesia, may prove to be a useful source of

INTRODUCTION 25

vanadinite, whilst smaller quantities are found in Western
Australia.

Zirconium. — In the form of zircon, large quantities
are available in the heavy sea sands of S. India and Ceylon.
As sylenite, it occurs in several localities in Scotland, Ireland,
Australia and Canada, but none of these sources have been
developed.

Copper. — In 1912, the world's copper consumption was
just over one million tons, of which Great Britain and the
Dominions provided one-tenth. 80 per cent, of the British
production is obtained from Queensland and New South
Wales, where in 1913 an output of nearly 50,000 tons of
metal was reached. Extensive deposits have been worked
in British Columbia and Ontario, whilst in South Africa,
the Cape Province, the Transvaal and Rhodesia offer fields
for further development. Chalcopyrite deposits are found
in Cornwall, Devon and N. Wales, and cupriferous pyrites in
Co. Wicklow, Ireland. It is stated that the deposits in
British New Guinea are to be developed on a large scale.

Lead and Zinc. — The world's lead production just
before the war exceeded 1,000,000 tons, and the zinc con-
sumption was much of the same order. In 1912, 25,500 tons
of lead ore and 11,700 tons of zinc ore were mined in the
British Isles, chiefly in Wales and W. England, and this was
exported for reduction. The chief source of lead and zinc
ores before the war was the Broken Hill area in New South
Wales, where 500,000 tons of ore, consisting of galena blende
mixtures containing pyrites with a gangue of garnet quartz
and rhodanite, were annually exported. Other Australian
deposits are found in W. Australia, Queensland, Tasmania
and New Zealand.

The chief lead and zinc ore producing area in Canada is
the Koolenay district, British Columbia, where over 86,000
tons of lead ore and 11,000 tons of zinc ore were raised in

I9I3-

It is claimed, according to prospectors, that the Rhodesian

Broken Hill deposits in North West Rhodesia exceed those

in Australia, and are capable of extensive development.

26 INDUSTRIAL ELECTROMETALLURGY

Deposits are also formed in Upper Burma, where the mines
were worked at a very early date by the Chinese for silver.

Tin. — The world's tin consumption is stated to be about
130,000 tons, of which the Malay States provide over 50,000.
In Cornwall, where tin has been mined for a very great
number of years, the annual production is still 5,000 tons,
and capable of further development and improvement.

Tin ores are also found in Burma, and less abundantly
in Australia, Tasmania, New South Wales and Western
Australia. In Africa a development of the Nigerian deposits
is to be expected.

Aluminium. — There is a marked scarcity of bauxite
deposits within the Commonwealth. Since the annual
world consumption of aluminium is in the neighbourhood
of 100,000 tons, and is rapidly rising, it is unfortunate that our
sole deposit is found in the Co. Antrim, Ireland, where the
annual output is equivalent to only 1,500 tons of metal.

Other sources from which the metal can be economically
obtained must be sought for, and production from these
alternative ores encouraged.

In India, British Guinea and the Malay States, extensive
deposits of laterite, a low grade bauxite rich in iron, appear
capable of economic development, whilst alunite, a hydrated
potassium aluminium sulphate, is found in New South Wales,
in India, and on Vancouver Island in Canada.

The possible utilization of the felspars must also be
considered as a future source of this metal.

INTRODUCTION 27

REFERENCE LITERATURE.

" Principles of applied Electrochemistry." Albnand.
" Electrometallurgy." Macmillan.
" Electrodeposition of Metals." G. Langbein.
'Practical Electrochemistry." B. Blount.
" A Textbook of Thermodynamics." J. R. Partington.
" Elektrochemie." S. S. Van Laar.
*' Elektrochemie." Le Blanc.
" Grundniss der Elektrochemie." H. Jahn.
" Theoretische Chemie." W. Nernst.
The Transactions of the American Electrochemical Society.
The Transactions of the Faraday Society.
Zeitschrift fiir Elektrochemie.

Section I.— ELECTROLYSIS IN AQUEOUS

SOLUTIONS

Copper.

The electrolytic refining of copper is by far the oldest and
largest of electrometallurgical processes, and has had a
remarkable development especially in America, where over
85 per cent, of the world's copper production is dealt with.
Of recent years some advance has been made on the electro-
lytic recovery of copper directly from the ore by leaching with
a suitable solvent and subsequent electrodeposition of the
metal.

The Electrolytic Recovery of Copper from its Ores, —

Early experiments such as those of Marchese and Nicola-
jew, made on coarse metal matte (Cu2S,Fe 2 S 3 ) and white
metal matte (Cu 2 S) obtained in the ordinary metallurgical
process, indicated that these sulphides, although easily cast
into anodes and of sufficient electrical conductivity for use
in electrolytic cells, using copper and ferric sulphate contain-
ing free sulphuric acid as electrolyte, were unsuitable for
this purpose owing to the rapid accumulation of impurities
in the electrolyte and the uneven solution of the anodes.
The liberation of sulphur —

Cu 2 S->CuS' +Cii->Cu+S"

which adhered to the anode caused the voltage to rise
above economical pressures, whilst only low current densities
could be used, 0*3 amp. per 100 sq. cms.

Borchers in 1908 conducted some experiments l at
Mansfield, in which the matte was further refined by blowing
in a Bessemer converter, then fusing the Cu^S now free from
the metalloids into anodes. An acid copper sulphate

ELECTROLYSIS IN AQUEOUS SOLUTIONS 29

electrolyte was used, and with a current density of 0*5 amp.
per 100 sq. cms. good deposits of pure copper were obtained.
Agitation of the electrolyte was found necessary to detach
the sulphur deposit from the anodes. For a short period
over 10 tons of copper were produced per week by this
method.

Attention was then directed to the method of leaching
out the copper from the crude ore or from concentrates.
Subsequent electrolytic deposition of the metal from the
electrolyte with insoluble anodes was employed, and the
spent electrolyte could then be returned to the leaching
vats.

(a) The Ferric Sulphate Process.

This method, originally suggested by Siemens-Halske,
utilized ferric sulphate as a solvent.

Oxidized copper ores such as the oxide or carbonate
can be crushed and leached directly. Sulphide ores con-
taining iron pyrites are roasted at a low temperature to
convert the iron sulphides into ferric oxide. The original
idea was to roast at such a temperature as to leave the
copper sulphide unchanged, but in practice the temperature
had to be elevated to 450 to 480 C. to ensure complete
conversion of the iron sulphide ; at this temperature most
of the copper is also oxidized. Dead roasting is to be avoided
owing to the possible formation of insoluble copper silicates
and iron copper oxide complexes as well as the possible loss
of silver. leaching with a 2 to 7 per cent, solution of ferric
sulphate is conducted in wooden vats, and solution of the
copper takes place according to the following equations : —

(1) Cu 2 S +Fe 2 (S0 4 ) 3 =CuS0 4 +2FeS0 4 +CuS

(2) 2CuS +2Fe 2 (S0 4 ) 3 +3O2 +2H 2 =2CuS0 4 +4FeS0 4

+2H 2 S0 4

(3) Cu 2 S+2Fe 2 (S0 4 ) 3 =2CuS0 4 +4FeS0 4 +S

(4) 3Cu 2 0+Fe 2 (S0 4 ) 3 ==3CuS0 4 +Fe 2 3 .

The ferric sulphate solution is produced by the atmo-
spheric oxidation of scrap iron dissolved in sulphuric
acid.

30 INDUSTRIAL ELECTROMETALLURGY

This process has been experimented with at the Ray
Mines, Arizona, at Cananea, and in a modified form at Rio
Tinto, Spain.

With ores containing 3 per cent, to 19 per cent, copper,
over 80 per cent, extraction can be obtained. Electrolysis
is conducted in a divided cell using thin sheet copper cathodes.
The problem of a suitable anode material for use in sulphate
baths has not yet been satisfactorily solved. Platinum is
ruled out on account of cost. Carbon and graphatized
carbon, although satisfactory in chloride electrolytes, are
rapidly destroyed by the oxygen evolution occurring in
sulphate solutions. Lead peroxide sheets (formed in situ
from lead sheet) have been successfully used, and on account
of their low cost are used for most technical operations.
Manganese oxide and fused magnetite (Fe 3 4 ) electrodes are
on the whole more satisfactory than lead, but more expensive.

As diaphragm for dividing the anode compartments
from the cathode, millboard asbestos is generally used.

Vertical electrodes are usually employed, although
horizontal ones have been suggested.

The cells are arranged in series, and the copper sulphate-
ferrous sulphate solution flows through the cathode com-
partments and returns through the anode chambers. Owing
to the deposition of copper on the cathodes during the flow
the electrolyte becomes specifically lighter, and conse-
quently enters each cathode chamber at the base and leaves
by the top. The reverse flow takes place in the anode
compartment (Fig. 1).

The ferrous sulphate produced from the interaction of
Cu 2 S and Cu 2 on ferric sulphate serves to depolarize the
anode according to the equations —

Cathode : CuS0 4 ->Cii +S0 4 "

Anode : SO" 4 +2FeS(V>Fe 2 (S0 4 ) 8
instead of the evolution of oxygen according to the equation —

2SO" 4 +2H 2 0->2H8S0 4 +0 2
The saving in the electrical energy required to bring about

ELECTROLYSIS IN AQUEOUS SOLUTIONS 31

the deposition of copper from a copper sulphate solution
with an anode depolarizer is very great and can be calculated
as follows : —

The minimum decomposition voltage of copper sulphate
with the deposition of copper at one electrode and the
liberation of oxygen under atmospheric pressure and with
no overvoltage at the other, can be calculated from the
thermal data of the reaction

t t

2CuS0 4 +2H 2 0->2Cu -r-HjjSOj-r-Os

requiring 56,300 calories per gramme atom of copper.

Fig. 1. — Arrangement of circulation, in cells for deposition of copper

from cnpric ferrous sulphate electrolytes.

A. Cathode compartment. B. Anode compartment

= I"22 VOltS.

The theoretical decomposition voltage is therefore
5 6,300 x 4-2 _
96,540 X 2

With ferrous sulphate as anodic depolarizer we have the
equation

CuS0 4 +2FeS0 4 -*.Cu-f-Fej.(S0 4 )3

requiring only 16,800 calories ; hence the requisite decom-
position voltage is

16,800 X 4'2

.0-36 volt.

96,540 X 2

The introduction of a diaphragm into the cell, however,

INDUSTRIAL ELECTROMETALLURGY

necessitates the use of a much greater externally impressed
electromotive force. In the experimental runs, between o*8
and i'8 volts were used with a current density of 02 amp.
per sq. dcm.

(b) The Sulphuric Acid Process.

The use of sulphuric acid as a leaching agent for copper
ores has advanced more rapidly than the ferric sulphate
process, and may be said to have outgrown the experimental
stage.

As in the ferric sulphate method oxidized ores can be
leached without any treatment, but sulphide ores first must
be roasted.

At the Chuquecamata mine in Chile, a large plant is in
the course of erection with a capacity of 335,000 pounds
of copper per day extracted from 10,000 tons of ore con-
taining brochantite (an oxy-sulphate of copper) averaging
about 2 per cent, copper. It is proposed to crush the ore
to pass a 0-25" mesh, to leach it with approximately 12 per
cent, sulphuric acid in concrete leaching vats lined with
mastic asphalt. It was found that the solution would be
efficiently filtered through cocoanut matting set between

resulting electrolyte : —

gms. per litre.

• 5044

371

0-07

0*06

Nil

CaO

080

MgO

3*32

Al 2 O s

i*6i

Na 2

21/60

K 8

5-00

S0 8

. 12275

a ..'

11-52

Free acid as H 2 S0 4

28-00

Solids on ignition

. 189/40

HNO3

ELECTROLYSIS IN AQUEOUS SOLUTIONS 33

It will be noted that both arsenic and antimony are
absent, both being very deleterious for copper deposition
(seep. 40).

The main objectionable impurity is the chloride, since
not only is part of the chlorine evolved at the anode during
electrolysis, but part is included in the deposited copper as
cuprous chloride ; it was therefore proposed to remove the
chlorides previous to electrolysis. This was effected by
agitating the solution with shot copper in revolving drums.
Cuprous chloride is formed according to the equation

2CuCl 2 +2Cu->2Cu 2 Cl 2

which can be filtered off, dried, fused with calcium chloride,
and reduced to metallic copper by smelting with coke. It
is proposed to use magnetite anodes 4 feet long, 5 inches
wide and 2 inches thick, with five to a vat, and the ordinary
sheet electrolytic copper anodes, 3 feet wide and 4 feet deep.
The spent electrolyte in the experimental plant contained
i*5 per cent, copper, and was returned at this stage to the
leaching vats. The average extraction was found to • be
close on 91 per cent.

I^aszczynski 3 obtained an efficiency of over 91 per cent,
with a current density of 0*5 to 1 amp. per sq. dcm., with an
electrolyte containing 3 per cent, of copper, obtained by this
leaching process.

At Butte Montana a 2 per cent, carbonate ore is treated
by leaching with 10 per cent, sulphuric acid after crushing
to J". Previous to electrolysis the electrolyte is heated by
steam to 6o° C. At a neighbouring plant electrolysis proceeds
with an agitated electrolyte using 1*4 amps, per sq. dcm.

Ricketts 4 cites a case of effective leaching with sulphuric
acid on an oxidized carbonate ore at Ajo, Arizona, the ore
being crushed to \" ; subsequent electrolysis using com-
posite coke-lead anodes gave a yield of 1 lb. of copper per
kw. hour (63 # 5 gms. per o # o8 kw. hr.).

Occasionally sulphur dioxide is injected into the anode

I*. 3

34 INDUSTRIAL ELECTROMETALLURGY

chamber of a divided cell to act as depolarizer, or the sulphur
dioxide may be made to agitate the liquid round the anode,
as suggested by Carmichael, in a simple cell. Hard rubber
tubes are most effective for conducting the S0 2 into the
electrolyte.

Sulphur dioxide is a more powerful oxygen depolarizer
than ferrous sulphate ; the critical decomposition voltage
for copper sulphate solution in a cejl anodically depolarized
with S0 2 is given from the thermochemical data —

CuS0 4 +S0 2 +2H 2 0-»Cu +2H2S0 4

with the evolution of 7,300 calories, or the theoretical
voltage is

—7300 x 4*2 „ 14 .

— ^ ~— = — 0-15 volt.

96,540 X 2

In other words, S0 2 should be able to precipitate copper from
a copper sulphate solution without the aid of any electrical
energy. This is actually the case, and forms the basis of the
Neill and Van Arsdale processes developed by Weidlein and
others for leaching and depositing from solutions. The
cycle of operations claimed for these processes is given in
the following equation : —

CuO+S0 2 =CuS0 3

Cupric sulphite is soluble in excess of sulphurous acid ; on
driving off the excess of sulphurous acid the cupric sulphite
is not deposited, but a red precipitate, Cu 2 S0 3 .CuS0 3 , is
formed. The cuprous cupric sulphite on heating under
pressure with sulphuric acid precipitates copper according
to the equation —

Cu 2 S0 3 .CuS0 8 +2H 2 S04=Cu+2CuS04+2S0 2 +2H 2

The cupric sulphate can similarly be converted into metallic
copper by the addition of sulphur dioxide and the inter-
mediary precipitation of the double sulphite.

In practice anodic depolarization is not complete,
usually only about 60 per cent. 6 and applied voltages from
o*2 to i*5 volts have been used.

At the International Copper Company's plant in Canada,

ELECTROLYSIS IN AQUEOUS SOLUTIONS 35

lead anodes separated i| inches from sheet copper cathodes
are used witlra copper content of 2*5 per cent, and 2*5 per
cent, sulphuric acid in the electrolyte. The applied voltage
is 1 "5 to pass 07 ampere per sq. dcm. with a current
efficiency of 90 per cent.

(d) Chloride Processes.

The original chloride process is that detailed by Hoepf ner
at work in Silesia. The finely crushed ore is leached with a
cupric chloride solution containing sodium or calcium
chloride heated to about 70 C. in wooden drums. I<ead
chloride would be removed on cooling, and the metalloids
and iron by lime. Owing to the presence of excess chlorine
ions the small amount of copper going into solution as
cuprous chloride becomes complex :

Cu+2Cl'^CuCl' 2

thus permitting of the preparation of relatively concentrated
solutions of copper in the cuprous state. The electrolyzer
consists of a series of divided cells with asbestos parch-
ment or perforated mica diaphragms. Carbon or graphi-
tized carbon can be used as anode material in chloride
solutions, although Hoepfner found these not sufficiently
refractory, and suggested the use of ferrosilicon. Sheet
copper cathodes are employed. In the cathode and anode
compartments the following reactions take place : —

t
Cathode : Cu 2 Cl 2 -> Cu+CuCl 2

Anode : Cu 2 Cl 2 +2Cl/->2CuCl ?

the cuprous chloride in the electrolyte acting as an anodic
depolarizer for the liberated chlorine. The catholyte and
anolyte consisting chiefly of cupric chloride, after passing
through the cells, are mixed and returned to the leaching pi ant.
The theoretical decomposition voltage obtained from
the thermochemical data —

2Cu+Cl 2 =Cu 2 Cl 2 +35,ooo calories,

is as follows : —

35,000 x 4*2 M «,

„**,,. = 1 '53 volts,
96>54o

36 INDUSTRIAL ELECTROMETALLURGY

the copper in this case being monovalent. The electrical
energy required to deposit a gramme molecule (63*5 gms. of
metal), viz. 96,540 x 1*53 watt sees. =0*043 kw. hour.

If cupric chloride were used as electrolyte the decom-
position voltage would be —

Cu + Cl 2 = CuCl 2 + 62,500 calories

,6^500 X 4 -a yolts

9 6 >54° x 2
= 1-35 volts,

the copper in this case being divalent. The electrical energy
required to deposit a gramme molecule (63-5 gms. of metal)
being 96,540x2x1*35 watt sees. =0*075 kw. hour.

There is, therefore, a distinct advantage in using cuprous
chloride instead of cupric chloride as electrolyte, although
since the decomposition voltage of the cuprous salt is higher
than that of the cupric, the energy gain is not quite that
to be expected by a change from the divalent to the mono-
valent state of the metallic ion, viz. double the output per
kw. hour.

The further advantage of anodic depolarization can be
calculated from the heat of reaction —

Cu+CuCl 2 ==Cu 2 Cl 2 +i9,400 calories

The theoretical voltage is therefore

19,400 X 4 '2 = Q . 8ol
96,540

or the minimum energy required to deposit a gramme

molecule (635 gms. of the metal) is 96,540 xo*84 watt sees.

or 0*022 kw. hour. The actual voltage required was said

to be o*6 to o*8 volt per cell. Early experiments in Saxony

(1892) proved unsuccessful, chiefly owing to difficulty of

leaching with cupric chloride according to the following

equation : —

2CuCl 2 +CU2S =2Cu 2 Cl 2 +S

Further investigation has shown that cupric sodium
chloride solution is a good leach for certain oxidized ores

ELECTROLYSIS IN AQUEOUS SOLUTIONS 37

such as CuSi0 8 found at Miami, Arizona, where a 2-hours*
leach with a 5 per cent, cupric chloride solution on a 3 to
5 per cent, copper ore ground to pass a 60-mesh sieve yielded
a 99 per cent, extraction.

At a current density of 11 to 1*3 amps per sq. dcm.,
i*o volt was required per cell. Greenawalt has used an
acid chloride leach with success, using S0 2 as acid. He
finds it desirable to roast ores containing much iron or
sulphides. A divided cell is not used in his process, and
with an applied voltage of 1*53 volts copper can be deposited
with an electrical eneTgy expenditure of 0*080 kw. hour
per gramme molecule (63*5 gms.) of copper.

Leaching with ferric chloride was the subject of a Belgian
patent (the Body process). The ferrous chloride formed
during the leaching acts as an anodic depolarizer, according
to the equation —

2FeCl 2 +Cl 2 =2FeCl 3

The theoretical voltage can be calculated from the heat of
reaction —

Cu4-2FeCl a =CuCl 2 +2FeCl 2 +7ooo calories ;
therefore, the minimum theoretical E.M.F. required

= 7000 __X4^ volt .

96»540 X 2

The Electrolytic Refining of Copper. — The raw

material for electro-refining is blister copper, obtained from
the ordinary smelting process. The metal analyzes some
98 per cent, copper, and is cast into bars some 3 feet long, 1*5
feet wide and 075 inch thick, and occasionally larger, to serve
as soluble anodes. Pure electrolytic copper sheet is used as
cathode material. The anodes and cathodes are suspended
alternately in wooden bitumastic or lead-lined vats carefully
insulated from the ground, some 2 inches af>art, although
with care the distance between the electrodes can be reduced
to as little as *5 inch, connected to copper bars which alter-
nate from tank to tank. About thirty pairs of electrodes are
used in each tank.

INDUSTRIAL ELECTROMETALLURGY

With a current density of from i'i to 2*2 amperes per
square dcm., the voltage loss per bath is approximately
0*2 to o*4 volt. A number of tanks are connected in series
sufficient to make up a ioo or 200 volt circuit. With a
current consumption of 2*2 amps, per sq. dcm. of cathode
surface and with 30 cathodes each of 100 sq. dcm. (3x1*5
X2 sq. ft.) a total current of over 5000 amps, per tank is
required. The modern tendency is to make the electrode
surface large and the tanks larger ; currents up to 15,000
amperes have been proposed. The average consumption
of current per pound of copper deposited is 0-166 kw. hour,
the ampere efficiency being 90 per cent. The electrolyte
consists essentially of an acid copper sulphate solution
containing 5 to 10 per cent, free sulphuric acid and 10 to

+..

4-..

Fig. 2. — General arrangement of electrolytic cells with plates in parallel.

15 per cent, copper sulphate which is continually circulated
from tank to tank to avoid stratification. The electrodes
are removed every three or four weeks. Since copper is
being dissolved at one electrode and deposited on the other,
theoretically no E.M.F. should be necessary to transfer the
copper, but, as has been already pointed out, a small P.D.
must always be applied to the electrodes, amounting to
from o*2 to 0*4 volt. Sixty per cent, of the fall in potential
across the bus bars falls on the ohmic resistance of the
electrolyte. In consequence the electrolyte in copper
refining gets warm owing to the energy absorbed, 60 per
cent, of 0*3 volt x 5000 amperes =900 watts per cell in the
above-mentioned case. Cooling by radiation normally
balances the heat energy supplied when the temperature
has risen to about 35 C, but owing to the high temperature
coefficient of the electrolyte it has been found economical

ELECTROLYSIS IN AQUEOUS SOLUTIONS 39

to still further heat it to about 55 £ with exhaust
steam.

An alternative electrode arrangement has been adopted
in some plants. The anode and cathode are end electrodes
in each vat, and the intermediate electrodes are bipolar,
copper being dissolved off one side and deposited on the
opposing face of the next electrode. The bipolar electrodes
are removed and the electrolytic deposit stripped off . When
the electrolyte becomes too contaminated for further use
the copper sulphate is partly removed by crystallization and
completely by the addition of scrap iron.

The Impurities present in Electrolytic Copper. —
Blister copper may vary widely in composition according
to the nature of the ore and the materials used in the
smelting process. Its copper content may fall as low as
91 per cent, or rise to over 98 per cent. The following
analyses indicate the usual impurities present and their
amounts : —

Copper . .

Arsenic . .

Antimony

Bismuth

Carbon

Iron

Iyead

Nickel

Silver

Gold

Zinc

Sulphur

Oxygen

The impurities in electrolytic copper should not exceed
o*i2 per cent. It is evident that it is more economical to
pay special attention to the preparation of high grade raw
blister copper than to attempt the electrolytic purification
of low grade anodes when such a high grade of purity is
demanded and can be obtained.

The impurities in refined copper may be due to a variety

2»

91-00

94-06

98*224

1*20

4*36

0-94

o'57

0*40

0'02

0*04

070

—

I'lO

OO3 2

trace

i'i6

OI3

0*02

1-14

o-37

0-28

1-09

—

30 oz. per ton

—

^q oz. per ton

1 '17

—

■ —

—

0'37

003

0*29

O'lO

0-05

40 INDUSTRIAL ELECTROMETALLURGY

of causes, 7 such as (i) the inclusion of the electrolyte between
the growing crystals on the cathode surface ; (2) electrolytic
deposition ; (3) mechanical contamination from the slimes.
These impurities adversely affect the quality of the deposited
copper by making it brittle and of low electrical conductivity,
or valuable by-products such as gold or more rarely platinum
and palladium may be removed in the cathode.

During the disintegration of the raw copper anode the
impurities in the copper either go into solution or fall to the
bottom as slimes. The impurities in solution may become
deposited in the cathode copper either by direct mechanical
occlusion or by electrolytic deposition whilst the slimes
become occluded by mechanical means.

Gold and silver are entirely eliminated as anode slimes,
none is f ound in solution. It is found that increasing current
density causes an increase in the gold and silver loss in the
cathode; this phenomenon is attributed to the greater
agitation of the electrolyte with the use of high current
densities increasing the quantity of suspended or " float "
slimes. Arsenic is present both in the slimes and in the
electrolyte ; it is probably deposited only by mechanical
occlusion of the electrolyte, since the quantity of arsenic
deposited varies only with the concentration of arsenic in
the electrolyte and is not affected by increased current
density.

Nickel which forms a continued mixture of solid solutions
with copper (Fig. 3) is probably electrically deposited, since
the difference of potential between the electrodes, viz. 0*4 volt,
would be sufficient to electrolytically deposit a solid solution
of copper and nickel containing very little nickel. The
decomposition potential of nickel sulphate is much higher
than that of copper, but the E.M.F. generated between
copper and copper nickel alloy containing but little nickel
can be made infinitely small. Oxygen is generally present
in deposited copper either as occluded gas or as cuprous
oxide in solicj solution. It is still a matter of speculation
as to the relative importance of these three factors, viz.
electrodeposition, inclusion of electrolyte and inclusion of

ELECTROLYSIS IN AQUEOUS SOLUTIONS 41

slime, on the amount and nature of the impurities in the
deposited copper. Generally it may be stated that an increase
of contamination with increasing current density and applied
voltage points to slime inclusion or electrolytic deposition,
whilst increasing contamination with increasing electrolytic
contamination points only to electrolytic inclusion.

^P Is

lickel

Copper

1 Alloys

•
LSbo C

;

\>>£j

*JV

i9ma

bji

\$^

l)0O
IOOO

'cyH

>a*°

Nh<**o'

*• 80 jo «*o £>•

7t> 60 W 100

100 AVom7 Co 1 00 A*bm % N i

Fig. 3. — Freezing'melting-pointfcurves'of nickel-copper alloys.

Electroplating with Copper. — In electroplating with
copper, pure electrolytic copper is usually employed as
anode material ; consequently, although no trouble is
occasioned by the presence of impurities in the metal or the
electrolyte, more attention has to be paid to the conditions
necessary for obtaining uniform, even, and compact deposits
of electrolytic copper.

The practice of violent agitation of the electrolyte or
movement of the cathode, e.g. for electrotype rolls, rotation
at a high speed is used to a greater extent than in copper
refining. Various electrolytic compositions aie in use, and
the current density is usually less than that employed in
refining processes.

The two most important electrolytes used for this work
are the acid sulphate and the cyanide bath.

42 INDUSTRIAL ELECTROMETALLURGY

•

Copper Sulphate Electrolytes. — Copper sulphate
electrolytes may vary in composition from a 5 per cent,
copper sulphate pentahydrate content to a saturated solution
with the addition of from o to 10 per cent, of free sulphuric
acid.

The current density usually employed varies from 1*5
to 2*5 amperes per square decimetre, although with high
rotational speeds for the cathode as suggested by S. Cowper
Coles up to 40 amperes per square decimetre can be employed.
The addition of iron salts to the electrolyte has frequently
been advocated. Mechanical burnishers such as agate
used by Elmore, sheepskin by Dumoulin, and glass beads
by Consiglio are frequently employed in addition to either
agitating the electrolyte or rotating the cathode.

Cyanide Solutions.— Copper plating is frequently used
for coating iron or steel as an intermediary film for nickel
plating. A thin film of copper can be deposited on the
iron by simple immersion of the iron in an acid copper
sulphate solution. Iron is more electropositive than copper,
and as has already been described the conditions necessary
for the deposition of one metal by another are directly
obtained by the. application of Nernst's hypothesis of
electrolytic solution pressures to metals.

If Efc and E Cu be the electrolytic' solution pressures of
the iron and copper respectively, and C F e, C Cii , be the con-
centration of ferrous iron and cuprion in the solution,
copper will be deposited at the expense of iron going into
solution, as long as

Epe ^ Ecu
Cf6 Cco

and the driving force or E.M.F. of the system will be

^log^-^log^ 1
2F ^ Cfc 2F Cca

The iron is, however, quickly entirely coated and deposition
ceases. In practice, however, u flashing " is liable to give
a porous and spongy deposit, and other electrolytes have to

ELECTROLYSIS IN AQUEOUS SOLUTIONS 43

be chosen. The choice of an electrolyte is limited to one
in which the E.M.F. of the hypothetical equation —

is reduced to zero or is practically negligible, i.e. copper must
exhibit no tendency to be deposited at the expense of the
iron. This can be obtained by the use of an alkaline electro-
lyte. In these solutions the concentration of cuprion is
depressed, making the right-hand term of the above equation
greater, and the iron is rendered passive, i.e. its apparent
electrolytic solution pressure E^ is lowered. The metal
becomes more noble or less easily attacked.

The following are some typical examples of such solu-
tions : —

(i) Copper salts and Cyanide.

Copper carbonate . . . . 100 gms./litre.
Potassium cyanide . . . . 200 gms./litre.

U.S. patent 129,124 describes a mixture —

Copper sulphate . . . . . . 100 gms.

Copper acetate . . . . . . 200 gms.

Potassium cyanide . . . . 150 gms.

Potassium carbonate . . . . 150 gms.

made up with 1 litre of water.

(ii) Copper salts, ammonia, and Cyanides. — Cowper Coles
describes the following solution as most effective : —

Copper sulphate 36 gms.

Ammonia (*88o) . . . . . . 26 gms.

Water . . . . . . . . 182 c.c.

mixed with —

Potassium cyanide 38 gms.

Water . . . . . . . . . . 148 c.c.

The solutions are mixed and made up to one litre.
Electrolysis is conducted with a current density of 0*4 to
0*5 ampere per sq. dcm.

INDUSTRIAL ELECTROMETALLURGY
Watt recommends the following solution : —

Copper sulphate
Water

230 gms.
1 litre.

to which ammonia (*88o) is added until the precipitate is
just redissolved. To this solution a strong solution of
potassium cyanide is added until the blue of the cupram-
monium cyanide is just changed to the lilac of the cupro-
cyanide complex. Electrolysis is best conducted at 6o° C.

(iii) Copper salts, Bisulphite, and Cyanides. — Baths con-
taining bisulphite have been recommended by Pfanhauser
and others. As typical of this class two may be mentioned —

(a) Na2S0 4 ..
NaHS0 3 ••
Copper potassium
Ammonia soda . .

> . . .
• . . .
cyanide

. . .

. 200 gms.
. 200 gms.
. 300 gms.
. 100 gms.

KCN

Water . . ■, .

. . .
. . .

10 gms.
10 litres.

(b) NaHS0 3 ••
Na 2 C0 3
Water

. . .

• . *

10 gms.

40 gms.

1 litre

mixed with —

Cu(CH 3 COO) 2 . .
NH 4 OH (-880) . .

40 gms.
14 gms.

with the addition of —

70 per cent. KCN
Water

. . .
. . .

56 gms.
1 litre.

Alkaline cyanide baths containing both tartrates and
thiosulphates have also been used with success. Other
alkaline baths which do not contain cyanide usually make
use of the solubility of copper tartrate in caustic soda.
Oxalate baths containing copper and ammonium oxalate
with free oxalic acid have been suggested by Classen,
Gauduin and others.

The Conditions necessary for Uniform Distribu-
tion of Copper. — We have already tabulated the various
types of electrolytes used in the electrodeposition of copper,

}

ELECTROLYSIS IN AQUEOUS SOLUTIONS 45

and have briefly referred to the practice of agitating the
electrolyte or rotating the cathode ; the exact mechanism
by which the cupric or cuprous ion is finally deposited out of
solution on the cathode, building up a solid coherent mass of
metal, is not yet clear, but the following considerations go
far to justify the combined use of many of the old recipes
and customs founded on experience or accidental discoveries,
and serve to sift out the worthless from an already extensive
list.

The Influence of Cathode Rotation or Electrolytic
Agitation on the Nature of the Deposit. — If two
unattackable electrodes, e.g. platinum, be immersed in a
solution of any salt, say copper sulphate in water, and an
externally impressed electromotive force be applied to the
electrodes, no electrolytic decomposition will take place if
the P.D. between the electrodes does not exceed the decom-
position potential of the salt. Although no visible electro-
lysis will take place, a small current will be observed to flow.
This " diffusion " current, owing to the different ionic
mobilities of the cuprion and the sulphation, effects a differ-
ence in concentration between the solutions surrounding each
electrode, as was first shown by Hittorf . After a short time
dynamic equilibrium sets in when the rate at which the
current tends to set up a difference in concentration is
exactly balanced by the rate at which the diffusion com-
pensates this change.

Sand 8 first pointed out the probable effect of raising
the applied electromotive force above the decomposition
potential of copper sulphate, but below that of sulphuric
acid.

At the cathode copper is deposited and the thin film of
electrolyte on the surface of the cathode has lost cuprion
to an extent which can be calculated from the current passing
through the cell. We have seen that the P.D. between a
metal and its solution is given by the relation —

E.M.F. = ^ log §»
2F 6 Ccu

46 INDUSTRIAL ELECTROMETALLURGY.

Consequently the back electromotive force of the cell
would rise to the value of the impressed voltage if the further
supply of copper ions to the electrolyte at the cathode
surface was not supplied by diffusion and by the electrical
migration of further cuprion —

— vc electrode + ve electrode

I<et Co be the concentration of the electrolyte at the
commencement of electrolysis. After a time / the concentra-
tion at a point x distant from the electrode becomes C», and

— is the concentration gradient from the electrode.

The quantity of salt diffusing per unit time across unit

dc
area with a gradient — is given by the equation —

where D is the diffusion constant. When 2=o, C=C
between x=*o and #= oo ; according to Fick's law the rate
at which the concentration alters is given by the equation —

fk — D—
dt dx*

or, on integration —

VttD J Vt

4D/

At the electrode itself, the concentration after a time interval
t becomes —

(!) C = C„ - aQ^/jL - Co - n 2 8 4 Q */I
The quantity of copper sulphate which has to be supplied

ELECTROLYSIS IN AQUEOUS SOLUTIONS 47

by diffusion (Q) is equal to the difference between that
deposited and that supplied by the ionic migration, or

q — __j — gm. equivalents,

* 96,540 96,540 s

where i is the current in amperes and n c the migration
constant of the cupric ion.

Substituting the value of Q in the above equation, we
obtain —

« *-*-&- V£

or, when C=o, i.e. at the moment when no cupric ions are
present in the layer of elctrolyte near to the electrode —

96,540 v 'V D

or- f=r C o 2D /9 6 '54Q>

(i-w c ) 2 tAri284/

When this time t has passed, further electrolysis can
only proceed by the discharge of hydrion at the electrode
at the higher potential difference necessary to decompose
sulphuric acid which is present in the electrolyte.

It is evijlent that under equilibrium conditions the
period of uniform deposition can be prolonged by starting
with a large value for C , the initial concentration, a high
diffusion constant and a high migration constant, which last
two factors can be sensibly increased by elevating the
temperature.

For the investigation on the beneficial effect of stirring
we can adopt the hypothesis of Noyes and Whitney 9 and of
Nernst, 10 applied to the solution of a solute in a solvent.
A solid in the course of solution, in their view, is to be
regarded as surrounded by a thin film of saturated solution
whence the salt diffuses into the less concentrated solution,
and the rate of solution is governed by the rate of passage
of salt from the saturated to the unsaturated solution.

48 INDUSTRIAL ELECTROMETALLURGY

If 8 be the thickness of the saturated film, then A the solution
constant per unit area is given by

A- D

According to Noyes and Whitney, when the rate of

i .. . dc
solution is —

| = A(C Mt .-C)

= g (Csat. - C)

where G^. is the saturation concentration and C the con-
centration of the surrounding solvent.

Applying this equation to the case of deposition under
consideration —

where Cq is the initial concentration of the solvent and C

dc
the concentration at the electrode. But -—, the rate of

deposition, is equal to Q, where —

q _ i(i — n c )
9 6 ,54<>

or Q=^(Co-C)

or C o"" C== f (3)

We have seen, however, that when no rotation or move-
ment of the electrolyte is' considered, from equation (1) —

o>- c -*y»

4=Vi>

82?T

or t — ~w

ELECTROLYSIS IN AQUEOUS SOLUTIONS 49

From this equation the conclusion can be drawn that

up to a time / = electrolysis can be continued with

4 D

the same current efficiency either with or without stirring

the electrolyte.

By rapid rotation of the electrode or agitation of the

electrolyte, the film thickness can be decreased and the

period of efficient deposition increased. A. Fischer n has

calculated from this equation the influence of rotation on

8, the film thickness in the case of a solution of copper

sulphate with the following results : —

Revolutions per minute

of stirrer.

Film thickness.

250

0*0635 mm *

800

0*0565 „

I IOO

00510 „

Thus, apart from the advantage to be derived from a
mechanical burnishing of the deposited metal by the circu-
lating electrolyte a distinct economy in time of deposition
by the use of higher current densities is effected.

The Influence of Simple or Complex Electrolytes
on the Nature of Deposit. — From the preceding con-
siderations we have noted that the electromotive force
between the copper and the surrounding electrolyte given

RT C

by the equation E.M.F. = — log ~ Cu is not constant, but

2F Ceo

may vary in a marked manner quite close to the electrode

due to the impoverishment of the electrolyte in cupric

ions by electrodeposition. The advantage of supplements

ing the supply of cupric ions normally migrating to the

cathode by agitation when high current densities have to be

employed is thus at once apparent, but the influence of the

concentration of the cupric ion in the solution on the nature

of the deposit is by no means so clear. v Since high current

densities and economical energy consumption are always

desirable, at first sight it might seem advantageous to lower

the E.M.F. between electrode and solution by increasing

the concentration of cupric ions. In practice, however,

1*. 4

50 INDUSTRIAL ELECTROMETALLURGY

for electroplating and electro-analysis, where a fine hard
coherent deposit is desired, quite the converse has been
found to hold ; namely, that a high cuprion concentration
round the electrode is not advisable in spite of the fact that
impoverishment may take place ; even if it occurs so
rapidly that, unless agitation be adopted, too low a con-
centration is arrived at.

We must assume that to obtain a hard deposit similar
to a worked metal we have to supply the extra energy
required to work the metal or burnish it electrically. In

RT F
other words, a limiting value is set to the term — log -^°

below which, although the metal is deposited, the deposit

occurs in a non-compact form. It must not be forgotten,

RT E
however, that if the term E == -=■ log — ^ be made too great

2F Ceo

by the depression of the cuprion concentration, disturbances

may occur due to the simultaneous deposition of other

ions, since the P.D. between solution and metal may

exceed the critical P.D. necessary to deposit the other

cation. The compactness of the deposited metal may be

attributed to the high potential gradient under which the

ion is deposited on the metal, or, on the other hand, to the

rapidity with which the ion changes into the metallic form

and loses its charge and water of hydration associated with

it, according to the following scheme : —

Cu(H 2 0),->Cu(H 2 0),r>Cu

There is strong evidence that the ions are hydrated in
solution, but we are not yet certain whether the loss of the
hydration water proceeds as rapidly as the loss of electric
charge.

Since in complex ion electrolytes the concentration of
the metallic ions is so low, the view is frequently held that
the electrolytic deposition of the metal is not a simple
direct discharge of the metallic ion as has been suggested,
but a secondary effect due to the discharge of other ions.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 51

For example, in potassium cuprocyanide, an anion complex,
the following equilibria are undoubtedly present : —

KaCu^CNJ^K'+zCuCCNJ'j
Cu(CN)' 2 ^Ctt+2(CN)'
KCN^K'+(CN)'
The value of

E =» ^ log g°? = _ 0658 volt.

* *~Cu

We obtain for a normal solution of cuprous salt and the
observed potential difference a concentration of cuproions
of io"~ 3 %. If cupric ions are assumed to be present, the
value of —

E = |£ log I* = - 0-329 volt,

which represents a concentration of io _48 w. cupric ions.

In the case of the copper ammonia or cupramine complex
as cation —

Cu(NH 3 )- 4 ^Cii+4(NH 3 )

A normal solution of cupric ions is reduced to io~~ 9 w. by the
formation of the cupramine complex, the P.D. rising from
— -0*329 to —0*0694 volt between metal and solution.

With these very small metal ion concentrations and
high electrolyte-electrode potential differences the discharge
of potassium ions, where the necessary P.D.

E= — log E -«
F * Ck-

must rise to +3*20 volts for a normal potassium ion solu-
tion, may possibly take place under working conditions ;
the potassium then secondarily deposits copper from the
solution according to the equation —

2K+K 2 Cu 2 (CN) 4 =2Cu+4KCN

Since equally good results are obtained with complex
metal electrolytes whether the metallic complex is an anion
or cation, Daniel's view 12 that the migration of the complex

52 INDUSTRIAL ELECTROMETALLURGY

away from the cathode at points of highest current density
results in an automatic readjustment of the current dis-
tribution can scarcely be correct.

The Influence of Various Addition Agents on the
Deposit. — In the previous section it has been pointed out
that in all probability the nature of the deposited copper is
greatly influenced by the conditions under which the dis-
charge of the cupric ion is brought about, the diagram-
matic scheme being —

Cu(R 2 0) tt ->Cu(H 2 0) tt -»Cu+2$

When the potential difference between solution and
electrode is great the cupric ion is discharged with great
speed at the electrode surface and the labile intermediate
complex has but a short existence. We can likewise in-
fluence this process of cathodic discharge not only by a
variation in the potential difference, but by influencing the
stability of the unstable metal hydrate which is a hypo-
thetical intermediary in the deposition. The nature of the
deposited copper can be made to vary so as to suit the
purpose for which it is intended. In electro-refining a
tolerable unif ormity of surface and density will suffice ; in
electroplating smoothness and hardness are desirable ; in
electroanalysis great smoothness and hardness are necessary ;
and for special work such as wire or tube construction
ductility* is necessary.

The Use of Colloids in Copper Deposition. — Vary-
ing the stability of the intermediary phase directly influences
the crystal size. The general practice is to add colloidal
substances to the electrolyte, when the following conditions
have to be observed. In acid solutions the added colloid
must be positively charged, i.e. it must tend to be codeposited
with the copper by electric endosmosis ; gelatine, glue and
tannic acid are colloids of this type. One part of glue in iooo
of acid copper sulphate electrolyte will give a fine-grained
tenacious deposit under the usual conditions of electrolysis.
The effect is more marked if the electrolyte be slightly
warmed to 25°-35° C. Such colloids act as protective

ELECTROLYSIS IN AQUEOUS SOLUTIONS 53

colloids to the labile metal hydrate and prevent the formation
of loose honeycomb structures or large crystals. As has
been pointed out by Bancroft, 18 the addition of colloids
tends to decrease the crystal size.

A second class of addition agent is to be found in the
form of salts which form insoluble colloidal hydrated
hydroxides in neutral and slightly alkaline solution, e.g.
tin and aluminium, and to a less extent, iron. In this case
the colloid is actually formed in the electrolyte round the
cathode, which tends to become alkaline during prolonged
electrolysis with indifferent agitation.

Successful experiments have also been made with non-col-
loidal addition substances, both reducing and oxidizing agents.

The Use of Reducing Agents in Copper Deposition.
— The deposition of copper is always attended with an
electrical energy loss due to the following reaction taking
place between the cathode material and the electrolyte. A
similar attack on the copper anode also occurs :

Cu+CuS0 4 ^Cu 2 S0 4
or Cu+Cu^Cu

As indicated by the arrows, the reaction is reversible, and
a rise in temperature shifts the equilibrium over to the
right-hand side of the equation. The small quantity of
cuprous ions which is present in solution tends to oxidize
on exposure to air in the acid electrolyte according to the
equation —

2Cu 2 S0 4 +2H 2 S0 4 +0 2 =4CuS0 4 +2H 2

The deposited copper can thus return again to its original
state in solution in the electrolyte. Since a very high
temperature or an electrolyte too concentrated in cupric
ions favours the formation of cuprous ions, these extremes
are to be avoided in copper deposition.

Reducing agents such as alcohol (3 to 5 per cent, in the
acid electrolyte), sugar, molasses, hydroxylamine and
pyrogallol have all been used. In pyrogallol and sugar
solutions, not only does the reducing power limit the influence

54 INDUSTRIAL ELECTROMETALLURGY

of the course of the reaction mentioned above, but sufficient
colloidal material is usually present to act as a protective
colloid.

The use of small quantities of chlorides up to '004 per
cent, have frequently been advocated ; not only does it
limit the concentration of the silver, antimony and bismuth
in the electrolyte (by the limited solubility of the chlorides
and oxyehlorides formed), but it appears to be beneficial
to the smoothness of the deposit ; the cuprous ions present
at the cathode will undoubtedly partially react with the
chloride to form the complex Cti + 2G'^CuCr. Deposition
of the cuprous ions will therefore take place from this
complex instead of from the simple ionized salt.

The Use of Oxidizing Agents in Copper Depositions.

— The use of oxidizing agents, usually nitric acid, is

practically confined to the electroanalysis of copper. We

have already noticed that under certain conditions of low

cupric ion concentration in sulphate electrolysis and high

current density, the difference of potential between electrode

RT E
and solution E = — log ^£ u may rise to such a point that

it approaches the potential difference required for evolution
of hydrogen. The rapid evolution of hydrogen causes the
deposited copper to assume a brown spongy appearance,
which has been attributed to the formation of an unstable
copper hydride or copper hydrogen solution. Nitric acid
acts as a cathodic depolarizer for the hydrogen, being reduced
to ammonia. 14

The Electrodeposition of Bronze and Brasses. —
Electrolytic depositions of " bronze " have been made, both
as true bronzes, namely alloys of copper and tin, and also of
brass or copper-zinc alloys. Nickel has also been substituted
for tin or zinc, and even arsenic has been used to obtain a
bronze-coloured deposit. The demand for bronze deposits
has been chiefly regulated rather by their colour and
appearance than by their actual composition.

The potential difference between a metal and a solution
containing its ions is given by the equation —

ELECTROLYSIS IN AQUEOUS SOLUTIONS 55

— RT , JSitn

where Ew=the electrolytic solution pressure of the metal,
and Cm the concentration of the n valent metallic ions in
the solution. For the simultaneous deposition of two
metals from an electrolyte containing both metals as ions
the following relation must hold : —

n{$ ° g Ctn x w 2 F ° g Cw 2

or Jiog— -l=log=-^

n 2 Cm 2 Ew 2

The following are the equilibrium potential differences
between metals and solutions containing the metallic ions
in normal concentration, taking the value

__RT 1 P. hydrogen a t i atmosphere

M „ M . ~" F . normal H ion solution

as zero : —

Copper
Zinc
Tin
Nickel

—0*329 volt (divalent ions only)
+0770 „

+0*192 „ (divalent ions only)
+0*228 „

Advantage is taken of the fact that the complexity of the
cuprous cyanide ion is much greater than that of zinc, nickel
or tin complex cyanides, and it is thus possible to raise the
deposition potential difference for copper to that point
where the other metal also begins to be deposited.

The degree of dissociation of the various complexes
varies with the temperature, and by electrodeposition from
a mixed complex electrolyte at various temperatures it
is possible to deposit either one metallic constituent or a
series of alloys.

The variation of the complexity is given by the cathode
potential at the various temperatures, and a few typical
examples are shown in the following curves : —

It will be noted that there is a very marked dependence
of the cathode potential on the current density. In these
cases since the ionic concentration of the metals in the

INDUSTRIAL ELECTROMETALLURGY

solutions is low and is removed rapidly by electrodeposition,
we must assume that the rate of reformation of metallic
ions by dissociation of the complex ion, e.g. Cu(CN)'g->
2CN'-|-Cu to arrive at equilibrium again is not instantaneous,
but proceeds with a slow reaction speed. The speed depends
in all probability on the complexity of the complex ion.
Tt will be noted that the silver complex breaks down most
rapidly, whilst the copper complex appears most stable.
The discharge potentials of the ions are of course

v^Agiefa — _&»& & ,& cr&r

1 a

1 M

- i ~

i--e a 6 — -i —

-- r Atf 2

IT /T

measured against their respective metals ; in the case of
deposition on to an alloy partial depolarization of the less
noble ion by the more noble constituent in the alloy will
take place, thus tending again to lower the divergence
between the two discharge potentials.

In the practice of alloy deposition we can control the
nature of the deposit by the followingindependent variables : —

(i) The complexity of the electrolyte affecting the ratio
C metal (i)
C metal (a)'

(2) The temperature affecting (1) and (3).

ELECTROLYSIS IN AQUEOUS SOLUTIONS 57

(3) The current densities affecting the velocities of

, . complex->ion (1)

reactions —. -. — ;-'

complex->ion (2)

(4) The composition of the deposit.

The following examples indicate some of the solutions
where the deposition potential of the two metallic ions is the
same : —

Copper-zinc.

(1) Copper sulphate 25 gms.

Zinc sulphate 29 gms.

Potassium cyanide to dissolve the precipitate.

• •

Water

(2) Copper acetate
Zinc sulphate
Potassium hydroxide
Potassium cyanide . .
Ammonia (*88o sp. gr.)
Water

Copper-tin.

(1) Cuprous chloride . .
Stannous chloride . .
Potassium cyanide • .
Potassium carbonate
Water

. .

. *

• •

a •

• .

i litre.

4'5 8 ms -
9-0 gms.

67*0 gms.

7'5 gras-
30*0 gms.

1 litre.

1-5 gms.

1 # 2 gms.

10*0 gms,

ioo'O gms,

1 litre.

(2) Copper phosphate 1 , ., « ,.

' Sodium pyrophosphate j saturated solution.

Sodium pyrophosphate 1
Stannous chloride J "

added to—

Sodium pyrophosphate
Water

(3) Copper sulphate
Stannic chloride
Potassium hydroxide
Water

(4) Copper sulphate
Stannous oxalate .
Ammonium oxalate
Oxalic acid
Water

• •

50 gms.
1 litre.

70 gms.

8 gms.
a small amount.

1 litre.

15-0 gms.

4*2 gms.
550 gms.

5-0 gms.

1 litre.

58 INDUSTRIAL ELECTROMETALLURGY

Zinc.

The conditions obtaining for the economical electro-
lytic recovery and refining of zinc are somewhat different
from those for copper.

The recovery of copper from its ores by purely thermal
processes is very economical, and it is only of recent years
that electrolytic methods have appeared feasible. In the
case of zinc, however, the recovery of the metal from the ore,
or more generally from the roasted ore, is one of great diffi-
culty. Not only are the retorts in which the reduction
takes place according to the equation

C+ZnO=Zn+CO

very fragile, owing to the extremely high temperature
necessary to bring about the distillation (over 1200 C), but
the condensation of the zinc vapour into a regulus cannot
be accomplished without the loss of blue powder (see p.
142).

Successful attempts have been made to use internal
electric heating for the reduction and distillation of the zinc ;
these will be dealt with in a later section.

With these disadvantages against reduction by means
of carbon, early attempts were made to utilize electrolytic
methods which may eventually entirely replace the early
thermal processes.

In the problem of electro-refining of zinc the case is not
comparable to that of copper. Not only is the demand for
very pure zinc limited except for shell manufacture, where
absolute uniformity in brass is necessary, but also the
purification of zinc by redistillation in vacuo of the crude
zinc obtained in the Belgian fuel furnaces is easily accom-
plished on account of the low boiling-point of the metal.
The cost of electrolytic refining of zinc is higher than that
of copper for reasons which will be stated. Thus the field
for electrolytic purification of zinc metal will probably
always be a limited one, but undoubtedly useful for working
up certain technical by-products, such as galvanizer's dross

ELECTROLYSIS IN AQUEOUS SOLUTIONS 59

(averaging 90 per cent, zinc and containing iron, tin and
lead) and zinc scum from the Parkes lead desilverizing
process (averaging 50 per cent, zinc and containing copper,
lead, silver and any gold that may be present). The alu-
minium amalgam modification of the Parkes process yields
a scum richer in zinc (70 to 80 per cent.) . It is also extremely
probable that the " blue powder " consisting chiefly of zinc
and zinc oxide, being the uncondensed portion of the zinc
regulus from the zinc furnace, could be more economically
disposed of electrolytically than by briquetting with carbon
and returning it in the furnace. With the growth of the
electric furnace production of zinc the quantity of " blue
powder " available for some such wet process will be
extremely large.

The Electrolytic Recovery of Zinc. — As in the case
of copper, many early attempts were made to use zinc ores
and roasted ores briquetted with coke as soluble anodes
in electrolytic cells, but were all unsuccessful.

Sulphuric Acid Processes. — Previous to the outbreak
of war, Germany controlled the greater part of the zinc ore
supply of the world mined in the Broken Hill area in
Australia. Their control was enforced by acquiring the
rights over the various flotation processes in operation to
concentrate the ore before shipping to Europe. This
ore is now available for the English market, and consists
of sulphides of zinc, lead, a little silver and gangue, being
essentially a blende galena complex.

The basis of all the sulphate electrolyte processes is the
primary roasting at a low red heat to convert the sulphides
into sulphates and oxides. The oxidized ore is thenleached
with dilute sulphuric acid and submitted to preliminary
purification before electrolysis.

The earliest electrolytic sulphate process was devised
by Siemens Halske and Laszczynski in Poland. Ten per
cent, sulphuric acid was used as a leaching agent. Lead
and most of the silver are removed as sulphate by filtration.
Small quantities of other metals, such as iron and copper
and soluble silica, are removed by fractional neutralization

60 INDUSTRIAL ELECTROMETALLURGY

with lime, followed by filtration. The presence of iron in
the ferric state is ensured by the addition of a little bleach-
ing powder. Frequently a final agitation with zinc oxide
or zinc dust is used before filtration. A nearly neutral
solution of zinc sulphate is thus obtained, which is circu-
lated through wooden electrolyzing vats containing sheet
lead anodes and thin electrolytic zinc cathodes. With a
P.D. of 3*8 volts per cell and a current density of i ampere
per sq. dcm., 3*4 kw. hours will deposit i kgm. of zinc,
showing a current efficiency of 80 per cent.

Modifications of this method are becoming increasingly
important for the electrolytic recovery of zinc. The follow-
ing difficulties were found to be inherent in the original
process : —

Anode Material. — The adoption of lead as anode material
was only made after extensive trials with other substances.

Carbon anodes, as has already been indicated, rapidly
deteriorate under the action of the oxygen evolved at the
anode. Early experiments by Ashcroft sought to eliminate
the oxygen evolution by the use of a divided cell, using ferrous
sulphate solution as a depolarizer, the ferrous sulphate
being oxidized to ferric sulphate ; cathodic reduction of the
ferric sulphate solution thus produced was accomplished
after most of the zinc had been deposited. In this case
ferric sulphate solution is used as leaching agent. Since
complete anodic depolarization is not required, and cathodic
reduction of the ferric sulphate proceeds practically quanti-
tatively, the extra number of ampere hours required to re-
oxidize the ferrous sulphate was practically compensated for
by the lower operating potential difference due to the anodic
depolarization.

Soluble iron anodes were also used, but it was found
in practice that the trouble associated with the use of dia-
phragms and the cost due to the loss of iron alone were
sufficient to make the process impracticable. Lead anodes
are now practically general, and usually prepared in situ .
from sheet lead, but as in the case of copper deposition carbon
protected by a thick deposit of lead peroxide, 1 * magnetite

ELECTROLYSIS IN AQUEOUS SOLUTIONS 61

and manganese oxides have experimentally shown better
results than lead sheet, so that there is reason to suppose
that sheet lead will be finally eliminated as soon as the
problems of uniformity of construction and low cost of manu-
facture have been solved.

The following table shows the excess voltage (or over-
potential) over and above the actual decomposition voltage to
be applied to ensure the liberation of oxygen at the anode : —

Nickel

Iron

Lead peroxide

Platinum (black)

Platinum (bright)

Graphite

Volts (overvoltagc
to oxygen).

. 0-05
. 0-24
. 0-28
. 024
0-44
. 0*40

Cathode Material. — These are usually thin strip electro-
lytic zinc about -fV to ¥ thick, separated by about 2" from
the lead anode. Difficulties are encountered in stripping
the zinc deposit from the plate, and it is generally found
necessary to form an artificial parting plane by slightly
coating the electrode before use. (Dilute rubber solution,
wax in alcohol, vaseline or glycerine are all effective.) Plates
hard to strip frequently strip on warming, but a certain
number have always to be melted up with the deposit.

A more serious difficulty is the corrosion occurring
at the union of the zinc plate with the copper connection
to the bus bar, and more especially at the surface of the
electrolyte (Fig. 5).

At the line of contact between the zinc and the electro-
lyte aa' the zinc is wetted with the spray, and since there is
no applied E.M.F. to keep the zinc from solution in these
acid drops, surface corrosion takes place. This is all the
more violent owing to the greater acidity of the electrolyte
'near the surface; the zinc sulphate solution, being denser
than the correspondingly concentrated sulphuric acid
solution formed by the electrolysis, always tends to gravitate
to the bottom of the electrolyzing vat and the acid to float

62 INDUSTRIAL ELECTROMETALLURGY

to the top, unless prevented from doing so by active circu-
lation. The corrosion is greatly assisted by the presence of
atmospheric oxygen, 16 and plates may be eaten through in
the course of a few hours. The usual method of prevention

Zinc

\jU

Fig. 5. — Surface corrosion of zinc cathode in zinc deposition.

is by the use of pure zinc cathodes which are not readily
attacked by acid, and if necessary by bitumastic paint to
just below the line of the electrolyte. Recently, the use of
aluminium plates as cathode material has been attended with
unqualified success.

Conditions for Deposition. — The conditions necessary
for obtaining uniform deposits from a sulphate electrolyte
have been the subject of many investigations, but the results
obtained are conflicting. In the case of copper deposition
the electrolytic potential of the metal Ecu referred to the
hydrogen electrode was —0*329 volt ; it is consequently
easier to deposit copper from an acid copper sulphate
solution than to liberate hydrogen. The electrolytic poten-
tial of zinc in a normal zinc ion solution is on the same
scale +0770 volt. It follows that if hydrogen and zinc
can be reversibly liberated or deposited at the anode of a
cell in an electrolyte containing normal zinc ion and normal
hydrion concentrations, hydrogen would be liberated before
any zinc could be deposited, until an excess anodic potential
of +0770 volt against the solution was applied above
that necessary to liberate the gas. Mylius and Fromm 17
also experimentally arrived at the conclusion that a high
concentration of zinc and a low acidity were most

ELECTROLYSIS IN AQUEOUS SOLUTIONS 63

desirable in an electrolyte. Further, it was found necessary
to work with a high current density.

The presence of basic salts is to be avoided owing to the
formation of a spongy deposit, and in practice it is necessary
to keep the electrolyte distinctly acid. The cause of spongy
deposition has been shown definitely to be due to the presence
of oxidizing impurities near the anode, and not to the forma-
tion of an unstable hydride of zinc, as was formerly con-
sidered. 18 Pring and Tainton 19 reinvestigated the problem,
and were surprised to find that with strongly acid solutions
and high current densities the deposition of zinc could be
carried out with a high efficiency, especially after the addition
of a small trace of colloidal material to the electrolyte.
Their process is now the basis of several semi-technical
deposition installations.

The electrolyte contains 150 gms. of sulphuric acid and
100 gms. of zinc sulphate per litre ; perforated sheet lead
anodes and zinc or aluminium cathodes, are used. The
potential difference over each vat is about 5 volts, and the
current density from 20 to 50 amperes per square decimetre.
An efficiency of 95 per cent, can be obtained at a tempera-
ture of 18 to 25 C. The zinc deposited by this method
from solutions containing manganese, lead, iron as grosser
impurities and small traces of other substances usually
obtained from roasted Broken Hill zinc concentrates, is
remarkably pure, averaging well over 99*80 per cent.

The curve in Fig. 6 represents the results obtained by
these authors, using 0*05 per cent, gum arabic as colloid
in an electrolyte containing 13 to 14 per cent, zinc sulphate
and 10 to 19 per cent, sulphuric acid with o*i per cent, of
iron.

The explanation of these results where the ratio of the
zinc deposited to the hydrogen liberated increases with
rising hydrion concentration in the electrolyte is far from
satisfactory.

As has already been pointed out in the introduction,
the overpotential necessary for hydrogen liberation at a
metallic surface varies with the nature of the metal. In the

6 4

INDUSTRIAL ELECTROMETALLURGY

case of zinc an applied E.M.F. of 070 to o*8o volt higher than
the reversible decomposition potential of the acid must be
applied to bring about the evolution of hydrogen. This
excess over the theoretical raises the critical potential
difference required to that necessary for the deposition
of zinc, which in a normal zinc ion solution is +0770 volt on
the hydrogen scale. In the neighbourhood of the electrode
under these conditions both ions are equally susceptible

100
93

I 80
I 75

)

So 3o Oo So 60 70 bo 9o

Amperes per »<j. dcm

Fig. 6. — Influence of current density on efficiency in deposition of zinc

to deposition, since the necessary deposition potential is
practically the same. The zinc ions have, however, a natural
preferment for deposition which may be explained on the
assumption that the velocity of deposition according to
some such scheme as follows : —

( A

I Zii(H 2 0) B

( B U C
|Zn(H 2 0)J Zn

is greater than that of the hydrogen deposition, which may
be depicted as

2H(H 8 0), J ~* 1 2H(H 2 0), J "* j H 2 (H 2 0), f ~* H 2

ELECTROLYSIS IN AQUEOUS SOLUTIONS 65

Not only have we other independent evidence that the
hydration numbers n and x are not the same for both ions,
but the second series of changes is bimolecular and not
an intermolecular change like the first ; both these factors
probably greatly influence the velocity of conversion.
Bennet and Thompson 20 believe that active hydrogen
(H as distinguished from H 2 or H") can deposit zinc from zinc
sulphate solution. If this assumption be correct a secondary
reaction between the hypothetical intermediary compound
2H(H 2 0)* and the zinc ions may occur according to the
equation —

Zn+2H-^Zn+2H

> »

Many investigators have accepted modifications of this
theory representing the change by the formation and decom-
position of unstable hydrides. The advantage of a high
current density is further emphasized by the consideration
that the resolution of the deposited zinc proceeds at a constant
rate for smooth deposits depending on the acidity of the
electrolyte, thus by increasing the rate of deposition the
apparent efficiency is also increased. Spongy surfaces
will naturally dissolve quicker than smooth ones, owing to
the greater area exposed to the solution.

The Use of Colloidal Addition Agents. — Pring and
Tainton recommend the use of colloids to ensure the deposi-
tion of the zinc and to eliminate impurities which are likely
to be deposited at a high working potential difference
between the electrodes. This point has already been
discussed in dealing with the deposition of copper. The usual
colloids employed are dextrin, gum tragacanth and gum
arabic of about 0*05 per cent, concentration.

Watts and Sharpe 21 suggest the use of 1 per cent, of
eikonogen, pyrogallol or j3-naphthol.

The Chloride Processes. — Hoepfner's original chloride
process was developed and is still worked by Messrs.
Brunner, Mond and Co., and is said to be at work at Duis-
berg and Fiirfurt in Germany, 22 but the use of blende in
preference to calamine as raw material has stimulated the

** 5

66 INDUSTRIAL ELECTROMETALLURGY

employment of sulphate electrolytes, more than the chloride
method.

A solution of zinc chloride is obtained by treating the
roasted zinc ore with calcium chloride in a carbonating
tower, when calcium carbonate is deposited according to
the equation —

CaCl 2 +ZnO +C0 2 -»ZnCl 2 +CaCO s

Alternatively the ore can be given a chloridizing roast with
salt. Iron and manganese are removed by the addition
of bleaching powder and a little alkali whilst a final filtration
over scrap zinc will deposit metals such as copper which may
be present. A 10 per cent, solution of zinc as chloride is
used as electrolyte, containing about 20 per cent, of sodium
chloride, a little free hydrochloric acid (o*i per cent.), and
gypsum. According to Foerster and Giinther, who carried
out experiments similar to those of Mylius and Fromm on
the sulphate solutions, the electrolyte must not be basic.

Operating with a diaphragm cell and a high current
density 3 to 4 amperes per sq. dcm. at 3-5 to 7 volts per cell,
good deposits of zinc analyzing 99 '97 per cent, may be
obtained provided that efficient circulation in the cathode
compartment is maintained. The use of revolving cathodes
possesses advantages for this process. The chlorine evolved
from the anode compartment where carbon anodes are used
can be used for preparing bleaching powder, for chlorination,
or may be compressed and liquefied. The diaphragms are
said to be of nitrated cellulose, but hydrated silica on
asbestos fibre has been stated to give, good results. It may
be noted that the addition of colloidal addition agents is
general practice. The cathodic current efficiency is stated
to be well over 94 per cent., whilst at the anode only 85
per cent, efficiency is obtained on the chlorine actually
collected.

The Electrolytic Refining of Zinc. — As has already
been mentioned, except in certain cases for the utilization
of by-products the method has but little commercial value.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 67

Richards successfully used galvanizer's dross as anode
material when cast with o-i per cent, aluminium. As
electrolyte he used 15 per cent, zinc sulphate hydrate, 17
per cent, commercial acetic acid, and 08 per cent, sodium
acetate. With zinc cathodes separated 4 cm. from the
anodes and a current density of 1 ampere per sq. dcm. at
a temperature of 30 to 32 C. and a voltage fall per cell of
1*25 volts, good deposits could be obtained provided that
air agitation and good circulation were employed. The
current efficiency varied between 80 and 100 per cent., and
the average analysis showed only 0*05 per cent, impurity.
The iron from the anode material was removed from the
electrolyte by the air agitation, followed by filtration of
the hydrated ferric hydroxide which was precipitated from
the acetate solution.

The Gold and Silver Anstalt at Hamburg attempted
the purification of the zinc scum obtained in the Parkes'
lead desilvering process, containing from 50 to 70 per cent,
lead, 10 to 50 per cent, zinc, and 5 per cent, copper, and
frequently up to 7 per cent, of silver and a little gold, and
o*2 per cent, aluminium.

They employed a zinc sulphate solution and either cast
anodes or granulated pieces lying on a horizontal carbon
anode. With a current of o*8 to 1 ampere per square dcm.
and at 1*3 volts per cell good coherent deposits of zinc could
be obtained, but the process did not prove commercially
successful.

A chloride process using zinc and magnesium chlorides
as electrolyte is said to be successful, in which the lead and
silver chlorides deposited in the sludge can be cupelled to
obtain the silver.

Electrogalvanizing. — Galvanizing is most commonly
accomplished by the hot galvanizing process, namely, by
cleaning the iron or steel plate, pickling it in acid, and
dipping it in a bath of molten zinc at a temperature of about
450 C. A superficial alloy is made with pure zinc on the
outside containing the compounds FeZny and FeZn 3 .

The formation of a film of iron-zinc alloy on the surface

68 INDUSTRIAL ELECTROMETALLURGY

may considerably lower the breaking strain of the thin
articles, such as hooks or cables, whilst the relatively high
temperatures employed (450 C.) may cause a lowering in
the tensile strength of steels due to this subsequent thermal
treatment. 28 Under these circumstances electrodeposition
from acid zinc sulphate electrolyte with lead peroxide anodes
can be feasibly employed.

The necessary conditions for deposition are identical
with those obtaining in the electrodeposition of zinc from
sulphate solutions, and have already been referred to.

Cadmium.

The electrochemical behaviour of cadmium is very
similar to that of zinc. Its deposition from solution, how-
ever, does not present such great difficulties as the former
metal, since its electrolytic potential on the hydrogen scale
is only +0*420 volt, whilst zinc has a value of +0770 volt.
Thus though from a solution containing both cadmium and
hydrogen ions hydrogen would be the first to be deposited,
yet, as was the case with zinc, the overpotential of hydrogen
against a cadmium cathode is very high, being +0*400
volt, making the conditions necessary for the deposition
of the metal with a high current density practically identical
with the former metal. Technically there is very little
demand for the pure metal, and the electrolytic recovery and
refining of the metal has not been accomplished on any scale ;
Brand 24 accomplished some large-scale experimental work
on purifying cast anodes of the following composition,
Cd 887 per cent., Zn 8*55 per cent., Pb 1*35 per cent., and
Cu 1*35 per cent. As electrolyte he followed the usual
practice of zinc refining, using a solution containing 10 per
cent, cadmium as cadmium sulphate and 5 per cent, free
sulphuric acid. His electrodes were spaced 5 cms. apart,
and successful deposition was accomplished with a current
density of 1*4 amps, per sq. dcm. The E.M.F. applied was
at first practically zero owing to the presence of the highly
electropositive zinc in the anode causing direct deposition

ELECTROLYSIS IN AQUEOUS SOLUTIONS 6q

of the cadmium. His final electromotive force was stated
to be only 0*048 volt per cell.

Electroplating with cadmium has a small technical
application. Under suitable conditions a soft white deposit
may be obtained which after buffing takes on a high polish
and resembles tin. Certain difficulties are inherent in electro-
plating articles with cadmium, which on deposition tends
to develop a macrocrystalline structure, a serious defect
when a smooth protective layer is desired. As in the case of
copper, this tendency can be corrected either by the addition
of suitable addition agents, usually colloidal, or by the
adoption of a complex electrolyte. In practice, practically
only complex electrolytes are employed. The usual
electrolyte is the complex cyanide formed by solution of
cadmium carbonate in a potassium cyanide solution.
Russell and Woolrich, 25 Fischer, 26 and Basset 27 all give the
composition of suitable plating baths. The electrolyte should
contain from 1 per cent, to 4 per cent, cadmium in the form
of cadmium carbonate dissolved in the minimum amount
of potassium cyanide necessary, and subsequently 5 per
cent, of potassium cyanide is added. Cadmium anodes are
usually employed, and uniform deposition is obtained at
a temperature of 40 C. with an applied E.M.F. of 3 volts.

Goi,d.

The electrolytic deposition of gold has been utilized
both as a means of obtaining the metal from a leaching
solution which has treated the ore and for the purpose of
plating less noble metals on an industrial scale.

The Electrolytic Recovery of Gold. — Gold generally
occurs in the free state as veins running through the auri-
ferous strata. When present in large quantities it can be
separated from the crushed ore by repeated washing with
water, the heavier gold particles being retained behind ;
frequently rough cloth or animal skins are used.

For poorer ores averaging only a few ounces to the
ton, chemical extraction methods are employed. The

70 INDUSTRIAL ELECTROMETALLURGY

earliest made use of mercury, as a solvent ; the gold amalga-
mates with mercury, which is subsequently removed and the
mercury recovered by distillation, leaving the gold.

There are several technical difficulties associated with
the ordinary amalgamation process. The mercury occasion-
ally " sickens " and becomes coated with a film of oxide,
hindering its coalescence and tending both to lessen its
power of amalgamation and to be carried away in the wash
water. Electrolytic reduction of the mercury oxide by
making it the cathode in an electrolytic cell or the addition
of a small quantity of sodium rectifies the tendency. The
gold- itself may be coated with the oxide or sulphide of some
other metal which may resist the amalgamating effect of
the mercury. Two other solvents are also employed for
the recovery of gold from its poorer ores, viz. free chlorine
and potassium cyanide.

In the process of chlorine extraction the ore is finely
crushed and extracted with an aqueous solution of chlorine
water prepared from bleaching powder and sulphuric acid.
This method of extraction is associated with the great
disadvantage that other metals are dissolved in addition
to the gold, and a very impure electrolyte results. More
common practice is the extraction by means of potassium
cyanide or potassium-sodium cyanide solution, in wooden
tanks with continuous agitation by compressed air. The
cyanide solutions possess the advantage that in dilute
solution the solution of gold is comparatively rapid whilst
other substances are relatively slowly attacked.

The dissolution of gold by potassium or sodium cyanide
solutions requires the presence of oxygen or an oxidizing
agent according to the equation 88 —

4Au+8KCN+2H 2 0+0 2 =4KA.u(CN) 2 +4KOH

Furthermore, the velocity of solution is greatly accelerated
by the use of a slightly alkaline medium. The addition
of sodium peroxide or potassium ferricyanide in small
quantities is said to increase the rate of solution to four or
five times the normal rate in the presence of air.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 71

The Electrolytic Deposition of Gold from Leach-
ing Solutions. — Deposition from a chloride or cyanide
solution can of course be accomplished chemically. The
zinc-lined boxes used for shipping the cyanide have been
used for this purpose, whilst for the chloride solutions ferrous
sulphate is generally employed. The Siemens-Halske pro-
cess for the recovery of gold possesses several advantages
over the chemical precipitation method. Very much weaker
solutions of cyanide can be used, down to as low as 0*05
per cent, cyanide, whilst for deposition by means of zinc
a solution at least ten times as strong is necessary. The
recovered gold is in a convenient form to handle, and the
electrolytic installation necessary is comparatively inex-
pensive to instal. Sheet lead cathodes and iron anodes
are employed. The anodes, 3 mm. thick, 2'i metres long,
and 0*9 metre wide, are enclosed in linen bags to prevent
the Prussian blue formed anodically by the action of the
cyanide electrolyte on the iron from contaminating the
electrolyte.

The current density employed is usually from 0*005 to
o-oi ampere per 100 sq. cm. with an applied voltage of
3 to 4 volts, decomposition of the cyanide taking place with
an E.M.F. above 5 volts. The electrolyte is slowly circulated
through large wooden vats 30 feet long by 6 feet by 6 feet,
which are divided into compartments so as to admit the
electrolyte at the top and exit at the bottom of the cell.

Auric cyanide rapidly dissolves in excess of cyanide to
form a practically colourless complex cyanide —

Au(CN) 3 +KCN$KAu(CN) 4

The complex auric cyanide is dissociated in solution into
potassion and the complex anion Au(CN) / 4, which is again
dissociated according to the equation —

Au(CN)'^tAu"+4CN'

As in the case of complex copper cyanides, a very uniform,
smooth and bright deposit of gold is obtained by this method.
The gold, averaging from 2 to 12 per cent, in weight of the
lead, is subsequently recovered by cupellation.

72 INDUSTRIAL ELECTROMETALLURGY

The disadvantages of the process are the difficulties
inherent in the use of iron as anode material, the con-
sumption of iron, and the contamination of the electrolyte
by Prussian blue. According to Blount, Andrioli employs
lead peroxide anodes and iron cathodes in a modification
of this process ; the lead peroxide anodes are said to be
unaffected by the electrolyte. Tin foil and carbon have
also been suggested. The gold deposited on the iron is
removed by immersion in a bath of molten lead and sub-
sequent cupellation. Keith suggests the co-deposition of
mercury and gold to facilitate precipitation.

In the Haycroft process an electrolyzed brine leach is
used, the chlorine being generated in situ by electrolysis
between a mercury cathode situated at the base of the
leaching chamber and carbon anodes suspended in the roof ;
the finely crushed ore in the brine is mechanically stirred
and the leaching vat kept warm. The gold is removed from
the ore partly by direct amalgamation and by electrolysis
of the auri-chloride formed by the action of the liberated
chlorine. The process does not seem to have passed the
experimental stage. Clancy 29 has conducted some promising
experiments on Haycroft's lines by using as electrolyte a
mixture of KCN, KI and KCNS and calcium cyanamide,
substituting the carbon anodes by the more refractory mag-
netite and using the iron leaching chamber as cathode.
Efficient solution of the gold is claimed due to the formation
of ICN at the anode. Cowper Coles has suggested the use
of a slowly rotating aluminium cathode for the deposition
of gold from a cyanide electrolyte. The gold deposit is
said to be easily detachable from the electrode surface, and
can be continuously removed in the form of a ribbon of thin
gold sheet.

The Electrodeposition and Refining of Gold. 80 — In
electroplating with gold, as in the case of the other metals
discussed, copper and zinc, advantage is again taken of
the uniformity and smoothness of deposits obtained by
the use of a complex electrolyte. For electrolytes contain-
ing less than o*i per cent, gold the temperature of deposition

ELECTROLYSIS IN AQUEOUS SOLUTIONS 73

should lie between 6o° and 70 ° C. Reddish matte deposits
are usually obtained. Cold electrolytes should contain
more than 0*4 per cent. gold. The more important complex
electrolytes used are the sulphocyanides, cyanide, ferro-
cyanide, and chloride; less important the phosphate, to-
gether with various electrolytes for producing coloured
deposits. 31

Cyanide Electrolytes. — The formation of a complex
cyanide on the addition of auric cyanide to a solution con-
taining excess of potassium cyanide takes place according
to the following equations : —

Au(CN) 3 +KCN^KAu(CN) 4

KAu(CN) 4 ^K+Au(CN)' 4

Au~+4(CN)'

When gold chloride is used as the source of the gold in the
electrolyte, primary decomposition takes place according
to the equation —

AuCl3+3KCN=Au(CN) 3 +3KCl

Fulminating gold, Au(NH 8 ) 2 (OH) 3 , is frequently formed as
an intermediary by precipitation with ammonia to avoid
the presence of chlorides in the electrolyte. Anodic solution
proceeds smoothly in potassium cyanide electrolytes, but
according to Jacobsen and Cohen, 32 in dilute sodium cyanide
solutions the metal is liable to become passive owing to the
formation of insoluble sodium aurous cyanide, NaAu(CN) 2 .
The following bath suggested by Roseleur may be taken as
typical of the cyanide electrolytes : —

Ten gms. of gold as chloride are dissolved in 250 c.c. of
water and mixed with 20 gms. of potassium cyanide (98-99
per cent, pure) in 750 c.c. of water. Langbein recommends
that this be boiled half an hour before use. Small current
densities, with anodes of pure gold sheet, are usually employed
from 012 to 0*41 ampere per 100 sq. cm., with a bath voltage
of from 27 to 4 volts. The optimum temperature of deposi-
tion lies between 50 and 6o° C. Dipping baths in which
deposition is brought about by the insertion of sheet copper

74 INDUSTRIAL ELECTROMETALLURGY.

or zinc usually contain less potassium cyanide, so as to in-
crease the concentration of gold ions in the solution.

Ferrocyanide Electrolytes. — The following reactions,
according to Beutel, 33 take place in the formation of the
potassium' auric cyanide complex from a gold salt and
potassium ferrocyanide : —

HCl.AuCl 3 +K 4 Fe(CN) fl +0 2 ->KAu(CN) 4 +KCl+KCN

+Fe 7 (CN) 18 +H 2

His numerical relationships are, however, so complicated as
to cast doubt upon this interpretation of the reactions taking
place. The ferrocyanide baths formerly had the advantage
over the cyanide electrolytes on account of their com-
parative cheapness and purity. With the modern methods
of cyanide preparation these advantages no longer exist.
They are not so poisonous as the cyanide baths, but on the
other hand do not dissolve the gold anode so readily and the
addition from time to time of auric chloride is necessary.
Pfanhauser 34 recommends the use of 15*9 gms. of auric
chloride, 90 gms. of ferrocyanide, with the addition of an #
equal amount of potassium carbonate per litre. The solution
is boiled and the ferric hydroxide precipitate is filtered off.
The temperatures and current densities are the same as those
employed for cyanide electrolytes.

Chloride Electrolytes. — This electrolyte, originally
suggested by Eisner ** and studied by Bottger and Neu-
mann, 36 was developed by Wohlwill, 87 and is the electrolyte
employed at Hamburg for refining gold by the N. Deutsche
Raflinerie. 38 Dr. Tuttle introduced the system with certain
improvements into the Philadelphia Mint, where a large
plant is now installed.

Crude gold containing both platinum and palladium is
used as anode material, and large thin sheet gold cathodes
are employed, the leads being of gold wire ; soldered joints
are avoided. The current density employed is, for the cathode
10 amperes per sq. dcm., and up to 30 amperes per sq. dcm.
for the anode ; the fall of potential over the bath is less than
1 volt. The electrolyte contains about 25 to 30 grammes of

ELECTROLYSIS IN AQUEOUS SOLUTIONS 75

gold per litre as chloride, and about 3 per cent, of free hydro-
chloric acid, the temperature being maintained at 50 to jo°.
The deposit of gold is uniformly pure and both adherent
and crystalline, especially when a little gelatin is added to
the bath. The solution contains the gold complex hydrogen
aurichloride, which undergoes partial ionization according
to the equation —

HAuCl^H- + AuCl'^H- +Au" +4CI'

It is important to have pure free hydrochloric acid in excess
in the electrolyte to ensure the uniform solution of the gold
anodes by the liberated chlorine. The primary formation
of some aurous chloride, AuCl, at the anode probably takes
place, with its subsequent decomposition into auric chloride
and gold, which is either redeposited on the anode or falls
as small crystals to the bottom of the cell —

3 AuCl =AuCl 3 +2Au

or is oxidized by the dissolved chlorine —

AuCl+Cl 2 =AuCl s

thus serving as an anodic depolarizer. A very small amount
diffuses into the bulk of the solution. At the cathode gold
will be deposited in excess of that demanded by the deposition
of trivalent gold due to the aurous ions present ; consequently
the weight of gold deposited is usually slightly more per
ampere-hour than would be obtained from a solution con-
taining only the trivalent gold ions. Platinum and palladium
are recovered from the electrolyte when sufficiently concen-
trated by the usual precipitation methods. They are not
cathodically deposited under the conditions of electrolysis.
Osmium, iridium and silver chloride are recovered in the
slimes. Over 76,000 ounces of gold per week are refined by
this process in New York and Philadelphia alone.

The anodic solution potential of gold in a chloride
solution is about E*=+i'i5 volts, indicating that the bulk
of the gold goes into solution in the trivalent state. On
raising the anode potential the gold is apt to become passive,
and chlorine will be liberated when the voltage has risen

76 INDUSTRIAL ELECTROMETALLURGY

to +173 volts. Addition of chlorine ions lessens the ten-
dency of the gold to become passive.

When relatively large amounts of silver are present in
the anodes the use of asymmetric alternating currents is
said to be attended with good results, preventing the silver
chloride from adhering to the anode and thus raising the
internal resistance of the bath. The use of bromide and
iodide baths has been the subject matter of a few early
patents.

Miscellaneous Electrolytes. — Withrow, 39 Perkin and
Preeble 40 obtained good deposits with Wallace and Smith's 41
modification of Von Ruolz's patent, which utilizes the
complex electrolyte formed on the addition of sodium
sulphide to a gold salt, or by the solution of auric sulphide
in excess of sodium sulphide —

AU2S3 +3Na 2 S^£2Na 3 AuS 3

The deposition of gold from this electrolyte, if similar to
that of antimony from its complex sulphide (see p. 90), is not
only due to the simple ionization of the salt according to
the following scheme : —

NaaAuSa^Na +AuS'" 3
AuS'" 8 ^Atr+3S"

but according to Ost and Klapproth, 42 the sodium sulphide
plays an important r61e —

NagS^Na+S"

At the cathode the discharged sodion reacts both with the
aurisulphide —

Na 3 AuS 3 +3Na =Au +3Na£S

and assists in the intermediary formation of aurosulphide,
according to the equation —

Na3AuS 3 +2Na =NaAuS +2Na 2 S

whilst at the anode the sulphur converts the monosulphide
into the yellow polysulphide —

Na2S+S=Na 2 S 2

ELECTROLYSIS IN AQUEOUS SOLUTIONS 77

The presence of excess of the poly sulphide is objectionable
if unattackable anodes are used as in electroanalysis, owing
to the solvent action of this salt on the deposited gold
according to the equation —

3Na 2 S 2 +2Au =2Na 3 AuS 3

The addition of sodium sulphate or potassium cyanide to act
as sulphur depolarizers have led to good results —

Na 2 S 2 +Na2S0 3 =Na 2 S 2 3 +Na 2 S
Na2S 2 +KCN=KCNS4-Na2S

At low current densities o # i to 0*3 ampere per sq. dcm.
at 6o° C, such electrolytes give excellent deposits.

Gold deposits can be tinted various colours by the
admixture with other elements such as arsenic, lead, or
more generally silver usually from cyanide baths. 43 Red
gold can be obtained by the addition of a small amount of
copper. One recommendation is to use both copper and
nickel in the electrolyte. 44

The Electrolytic Parting of Gold and Silver.—
Not only does natural gold contain a certain amount of
silver, from 5 to 50 per cent., but the silver slimes obtained
in copper refining (see p. 40) also contain gold ; according
to Pring the average composition of silver slime is 15*3
per cent, copper, 45*5 per cent, silver, and n per cent. gold.
The problem of parting gold from silver is therefore an
important one in both these industries.

The silver slimes from the copper deposition tanks are
washed, mixed with a small quantity of lead, and cupelled
to dore bars, the arsenic and other impurities being volatilized
during the process of cupellation.

The chemical process of parting by enrichment with
silver until the alloy contains approximately only 20 per
cent, gold with subsequent solution of the silver in nitric
or sulphuric acid leaving the gold unattacked is being sup-
planted by the electrolytic method introduced by Moebius
at the Deutsche Gold und Silber Scheide Anstalt at Frank-
furt a. M., and is at work in mints at New York and Phila-
delphia.

INDUSTRIAL ELECTROMETALLURGY

The electrolysis is conducted in earthenware or wooden
tanks, 2 ft. 6 in. deep and 3 ft. long, containing as electro-
lyte a mixture of nitric acid o - i to 1 per cent, and 2 to 4 per
cent, silver nitrate 45 usually with a varying amount of
copper nitrate when copper slimes are used. The dore metal
anodes, J in. by 5 in. by 12 in. in size, enclosed in canvas
or filter cloth bags, are separated about 6 in. from one another.
Silver foil cathodes are inserted 3 in. distant from each
anode. The silver is deposited at a high current density,
usually from 2-3-5 amperes per 100 sq. cm. at 14-17 volts,
to avoid interest charges on the silver. The loose feathery
crystals which have to be mechanically detached from the
electrodes are swept into canvas bags placed at the bottom

Fig. 6a.

Mechanical scrapers for the n

Dvftl of deposits of silver crystals.

of the vats. The mechanical scrapers usually employed,
which also serve to agitate the electrolyte, are of wood and
are of one of two forms. In the early form a wooden fork,
the prongs of which scraped the two surfaces of the cathode
plate, was suspended by a roller on a rail placed above each
cathode and caused to run backwards and forwards, scraping
off the crystals in its passage. A simplification introduced
in America consists of a fork suspended some distance above
the cathode and caused to oscillate backwards and forwards
about its point of suspension (Fig. 6b).

The silver crystals, which should contain no copper
provided that the acidity of the bath is kept high and the
current density employed not too great, are removed on
the trays, allowed to drain, washed and melted into ingots.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 79

The black pulverent anode slime, if washed and melted,
consists of practically pure gold, but is liable to contain
traces of lead or bismuth, or small pieces of the anode which
have dropped off during the process of dissolution may
contaminate the gold with silver and copper. These can
be removed by treatment with acid. The slimes thus
treated are cast into anodes and electrolytically refined for
gold. Modifications of the plant have been suggested with
a view to the elimination of the wooden scrapers, such as
the employment of a moving silver band as cathode. It is
placed at the bottom of the vat with a number of horizontal
anodes separated from it by canvas diaphragms placed
above. The process is in use at Monterey in Mexico. The
crystals deposited on the moving cathode are removed by
scraping and elevated out of the bath by another travelling
band.

At Balbach, U.S.A., Thum's modification of the Moebius
plant is worked with success. Horizontal anodes separated
by cloth diaphragms are employed as in the Mexican works,
but the travelling silver band cathode is replaced by graphite
block cathodes on which the silver crystals are deposited.
A slightly lower current density is employed, viz. r8-2
amperes per 100 sq. cms. at a higher voltage, 3*5 volts, owing
to greater distance between the electrodes and the inter-
position of the slimes. Mechanical agitation is dispensed
with, but the crystals are pressed down from time to time
to the bottom of the vat.

The conditions necessary for the separation of silver
without any copper in the electrolytic parting of gold and
silver are in practice very simple, viz. a high acidity and a low
current density. As, however, the metals locked up in the
vats are a great deal more valuable than copper, low current
densities are even more economically unsound than in
copper deposition, and in practice must be maintained as
high as possible. From 2 to 3 amperes per 100 sq. cms. with
an E.M.F. of 1*2 to 2 volts per cell are usually employed,
although in certain cases up to 6 amperes per 100 sq. cms.
have been used, the current density' being decreased as the

80 INDUSTRIAL ELECTROMETALLURGY

concentration of the copper salts increases. In a solution con-
taining normal cupric ion and normal silver ion the discharge
potentials of the copper and silver are —0*324 and —0771
volt respectively, there being a difference between the
two discharge potentials of nearly 0*5 volt. The decom-
position potential voltage of a normal silver nitrate solution
is about 070 volt, and since in practice the electrolytes
used are considerably weaker than normal, being approxi-
mately between o*i and 1 normal in respect to the silver,
this minimum decomposition voltage is therefore slightly
higher than 070 volt, and can be raised nearly 0*5 volt
without any copper commencing to be deposited. The
usual operating voltage lies between 1*28 and 1*35 volts.
In the processes carried out in the mints where the anodes
contain over 30 per cent, gold, no diaphragms are used, but
the vats are run at a low current density of o*8 ampere
per sq. dcm., attention is paid to obtaining an adherent
deposition, while the gold remains behind as an anode
skeleton. The addition of free nitric acid is necessary, up
to 1 per cent, acid, to neutralize any ammonia which may
be formed by the possible reduction of the nitrate ion taking
place at the cathode. The presence of even small quan-
tities of basic salts results in a formation of a spongy deposit.
Occasionally the silver crystals which are deposited are not
white, but dulled due to the formation of an unstable oxide ;
the addition of a small quantity of alcohol corrects this
tendency. Large crystals can be reduced in size by the
addition of 1 part in 10,000 of gelatine, but the addition
of gelatine must be made every day, as it is partly destroyed
by the nitric acid and partly removed in the deposited
silver. When the electrolyte has become rich in copper
salts (0*4 per cent.), the silver in the spent electrolyte can
be recovered by the addition of copper or precipitation
as chloride. Subsequent removal of the copper by electro-
lysis or chemical deposition with iron is usually employed.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 81

Shaver.

The electrolytic recovery of silver from its ores by the
application of the methods of electrochemical deposition
from one of the usual leaching agents employed in the
wet processes of silver extraction does not seem to have
received any attention, chemical precipitation by means
of scrap iron or copper being usually employed. Present
day economic conditions have shown that the electrolytic
winning of copper may be remunerative in certain localities,
and the electrolytic recovery of silver would probably be
even more favourable. As in the case of gold a cyanide
leach would probably offer several advantages. •

The electrolytic refining of silver is now practised
extensively, utilizing crude silver containing gold, copper
and lead together with many minor impurities. The
Pennsylvanian Lead Co. at Pittsburg use crude silver
anodes containing 2 per cent, lead, bismuth and copper,
whilst the New York and Philadelphia refineries use 30 per
cent, gold, 60 'per cent, silver, and 10 per cent, base metal
as anode material. Electrolytic refining could possibly
be substituted for cupellation of the zinc-lead-copper com-
plexes obtained in the various processes for removing silver
from lead. The parting of gold and silver as well as the
practical conditions to be observed in the refining of silver
from nitrate electrolytes have already been discussed.

Electroplating with Silver. — The nitrate electrolyte
is unsuitable for electroplating ; the deposit is macrocrystal-
line and spongy, probably owing to the formation of a sub-
oxide 46 or due to the absorption of oxygen. 47 The deposit
can be improved by rapid agitation or rotation of the cathode
as shown by Sand 48 and Snowden, 49 by the addition of
alcohol as suggested by Kiister, 60 or by the addition of
small quantities of colloids such as gelatine. These improved
silver deposits, although sufficiently good for silver refining
purposes or even for electroanalysis, are not suitable for
plating.

Cyanide Electrolytes. — The cyanide complex silver

Iy. 6

82 INDUSTRIAL ELECTROMETALLURGY

electrolyte is probably, in common with those of copper and
gold, the most suitable electrolyte for silver deposition.
Dissociation in the electrolyte proceeds according to the
following equations : —

AgN0 3 +KCN=AgCN+KNO s

The precipitate of silver cyanide is soluble in excess of
cyanide to form the soluble potassium silver cyanide which
is dissociated —

Ag(CN) +KCN < _KAg(CN) 2

KAg(CN)2<;K+Ag(CN)' 2
Ag(CN)'2$Ag+2(CN)'

The presence of the salt formed due to the decomposition of
the silver salt by the potassium cyanide has a considerable
influence on the nature of the deposit, the nitrate, chloride,
oxide, used originally by A. & H. Elkington in Sheffield in
1840, and carbonate of silver have all been advocated, whilst
other investigators insist on the primary separation of the
insoluble silver cyanide from the soluble salt making up the
electrolyte. Langbein advocates the use of the chloride,
but states that beyond certain limits the presence of chlorides
is apt to give the deposit a coarse structure. 61 Pfaunhauser
uses 10 gms. of silver as chloride and 20 gms. of potassium
cyanide per litre. With electrodes 10 cms. apart and a current
density of 0*3 ampere per 100 sq. cms. the potential drop
across the bath being about 1*25, a good deposit is obtained.
For heavier coats he suggests 25 gms. of silver as chloride,
with 40 gms. potassium cyanide per litre and the same
current density.

" Striking " baths for giving a preliminary thin coat for
certain work such as steel are generally very weak in silver.
A good electrolyte contains about 1*5 gms. of silver with
70 gms. of potassium cyanide per litre. A high current
density should be employed to ensure a brisk evolution of
cathode hydrogen. Foerster and Namias 62 advocated double
cyanide baths without the presence of any neutral salt. The
former suggests 25 gms. of silver cyanide and 25 gms. pure

ELECTROLYSIS IN AQUEOUS SOLUTIONS 83

potassium cyanide per litre, using a current density of 0*3
ampere per 100 sq. cms., with a P.D. of 1 volt.

The use of addition agents to cyanide electrolytes for
obtaining bright instead of matte deposits is very usual,
especially for plating articles which cannot easily be bur-
nished. Carbon disulphide has been used as an addition
agent since 1847. The quantity added should not exceed
2*5 parts per 10,000 ; agitation of the bath should be avoided, '
and the current density should be a little higher than normal.
Other but less effective addition agents have been sug-
gested from time to time ; amongst the more important
may be mentioned iodine or iodine and guttapercha in
chloroform, or a mixture of sulphur and collodion. The
use of a suspension of silver sulphide has also been suggested.
The use of these addition agents as brighteners appears to be
a particular case of the action of protective colloids such as
glue, linseed oil, mucilage or gelatine.

Miscellaneous Electrolytes. — Some of the earlier
experimenters advocate the use of ferrocyanide electrolytes.
Eisner 63 dissolved 7 gms. of silver in a solution of 8*4 gms.
of potassium ferrocyanide, 56 gms. of '88o ammonia, and 1
litre of water. These solutions have not been extensively
used, as they do not dissolve the silver anodes in a regular
manner.

Krutwig 64 claimed that silver could be deposited from
a silver hydroxide ammonia electrolyte provided that rapid
agitation of the electrolyte was ensured. The presence of a
reducing agent such as sulphurous acid or sodium thiosul-
phate is necessary. Various organic acids such as the lac-
tates, acetates, citrates have been the subject of patents,
but are not so efficient as the cyanide electrolytes already
discussed.

IvBAD.

The Electrolytic Recovery and Refining of Lead.—

The common lead ores consist of lead zinc sulphide com-
plexes containing varying amounts of gold and silver. In
the usual thermal treatment the sulphide ores are first

8 4

INDUSTRIAL ELECTROMETALLURGY

roasted. During the process of roasting two series of re-
actions proceed simultaneously according to the equations —

2PbS+ 3 2 =2PbO+2S0 2 l Roastin _ orocesses
PbS+20 2 =PbS0 4 } Koasung processes.

PbS+PbS0 4 =2Pb+2S0 2 1 Redllction oroC esses
PbS +2PbO =3Pb +S0 2 J Reductl0n Processes.

• If the general procedure of adding lime be followed a further
side reaction takes place —

PbS0 4 +CaO -± CaS0 4 +PbO

This roasted ore, <

containing varying

amounts of PbO,

PbS0 4) and lead, is then reduced in a blast furnace by means

of coke. The molten lead separates to the bottom, leaving

on top a mixture of lead, iron, and copper sulphide. The

crude lead so separated has approximately the following

composition : —

• • • •

• 98*3

• •

. o-i86

• • <

. 0720

• • i

0006

• •' <

. 0005

Ag .

• ■ <

. 0*141

■ • <

0001

• • <

. 0003

• ■ i

. 00023

• •

. 0*0002

• • <

. trace

Frequently also a small quantity of gold. It is then sub-
mitted to refining processes which will be described later.
The purely thermal process of roasting and reduction to
obtain crude lead is an economical one, since the heats of
formation of the oxide and sulphide are low, permitting
of easy reduction, and the low melting-point of the metal
ensures an easy removal from the furnace. Any electrolytic
treatment of the ore that could compete with this process
would be one in which the direct production of the pure
metal and the other by-product sulphur, either as such or as
hydrogen sulphide or sulphur dioxide, was ensured with the

ELECTROLYSIS IN AQUEOUS SOLUTIONS 85

minimum expenditure of electrical energy ; at the same time
permitting of the extraction of the valuable impurities in
the ore by some simple process.

It has generally been assumed that the low cost of
purely thermal processes would prevent the development
of any electrolytic process on a technical scale. The follow-
ing calculation will show, however, that if some such process
could be developed, the economic aspect of the question
is entirely in its favour : —

One ampere second will deposit 1*072 mgms. of lead, hence
a metric ton (1000 kgms.) will require 277 kiloampere hours.
Lead sulphide can be decomposed with an applied E.M.F.
of about ri volts, or 1 metric ton of lead could be deposited
by 300 kilowatt hours. With a kilowatt hour costing as
much as o^d. this only entails an expenditure of 12s. 6d.,
whilst the estimated cost of thermally refining crude lead
alone exceeds 25s.

Betts and Valentine 66 have made several experiments
on the electrolysis of lead sulphide dissolved in molten
lead chloride. They state that a good deposition of molten
lead can be obtained below a red heat with an applied E.M.F.
of ro to 1 25 volts. The presence of impurities in the
galena, however, has prevented this process from being
developed on an industrial scale. Anderson 66 attempted
unsuccessfully to electrolytically reduce galena in a fluosilicate
solution. In the Salom process worked at Niagara, lead
sulphide finely ground was admitted into a lead chamber
serving as cathode and container, with a 10 per cent, sulphuric
acid electrolyte. At a voltage of 25 to 2*9 volts per cell
a current efficiency of 70 per cent, was attained, the sulphide
being cathodically reduced to spongy lead and H2S.

Scarcely any attempts have been made to work up the
roasted ore electrolytically. The problem is analogous to
the recovery of lead from the lead sulphate scrapings ob-
tained in the lead chambers of sulphuric acid works. Lead
sulphate is soluble in sodium acetate and caustic soda ; from
both these electrolytes good deposits of lead may be obtained.
In the case of the roasted ore which contains, in addition

86 INDUSTRIAL ELECTROMETALLURGY

to the lead sulphate, both metallic lead and lead oxide, the
possibility of casting it directly into anodes presents itself.
Burleigh 67 suggested the solution of the roasted ore in hot
concentrated soda, where deposition of lead could be ob-
tained with an impressed E.M.F. of 17 volts.

The Refining of Lead. — Although not much progress
can be recorded in the electrolytic process for the recovery of
lead several schemes have been suggested for refining thecrude
lead electrolytically, and of recent years various improve-
ments have so modified the process that it is now much more
economical than either the PattinSon or Parkes refining
processes. In the Pattinson process the crude lead is sub-
jected to an oxidizing melt. The bulk of the zinc, iron, and
nickel is removed by steam injection, whilst the tin, arsenic,
and antimony are removed by introducing air, forming stan-
nate, arsenate, and antimonate of lead, which come to the
surface and are removed. On fractional crystallization
the first fraction consists of a copper-lead alloy which con-
tains the rest of the nickel, cobalt, sulphur, and arsenic ;
removal of the bismuth is never complete. In the Parkes
desilverizing process zinc is added to the partially purified
molten lead, when an alloy of gold and silver is formed which
solidifies on the surface of the molten lead. The solidified
alloy is removed and the zinc removed by distillation.
During the process of distillation a small quantity of silver
is also lost up to i\ per cent. For leads very rich in silver
(over 12 oz. per ton) the whole of the lead can be removed as
litharge by an air blast, leaving the silver and gold behind
on the cupel, the oxide lead being then again reduced to
metal.

The earliest electrolytic process for refining lead was
that of Keith. Crude lead anodes in muslin bags to retain
the slimes were used in an electrolyte of lead acetate or
lead sulphate dissolved in sodium acetate. Lead was
deposited on the thin sheet lead cathodes as small crystals,
which fell to the bottom of the cell and were removed and
fused together. The electrolyte contained 20 gms. of lead
sulphate and 150 gms. of sodium acetate per litre. The

ELECTROLYSIS IN AQUEOUS SOLUTIONS 87

current density employed varied from 0*2 to 0*35 ampere
per 100 sq. cms. at 0*4 to 05 volt. Tommasi employed a
rotating cathode in the form of an aluminium bronze disc of
3 metres diameter mounted on a horizontal axis performing
2 rotations per minute. A current density of 72 amperes
per 100 sq. cms. could be employed. By means of scrapers
on each side of the disc the lead crystals could be removed
on to a sieve conve}'or to be drained, washed and fused with
a little charcoal.

The Betts 68 process is in use at Trail, B.C., near Chicago,
and at Newcastle-on-Tyne, England, and may be said to be
the most successful of electrolytic lead- refining processes.
Crude lead is melted and cast into anodes about 75 by 75
by 2 cms. extending to 3*8 cms. at the top in size. Each
anode is cast with lugs and weighs about 170 kilos, being
separated from the next anode by a distance of 11*3 cms.
The cathodes are refined sheet lead not over 12 cms. thick
when finished. The electrolytic tanks are 6 feet long,
2 feet 6 inches wide, and 3 feet 6 inches deep, made of wood
lined with bitumen, $nd hold 22 anodes and 21 cathodes
each. The current density employed varies from 0*9 to 2*2
amperes per 100 sq. cms., and the applied E.M.F. from 0*15
to 0*42 volt, the E.M.F. gradually rising as the slime adheres
to the anodes. Even when the anodes are nearly com-
pletely dissolved they still retain their original form. The
electrolyte consists essentially of a solution of lead silico-
fluoride in free fluosilicic acid, first suggested by I^eucks. 60
Thirty-five per cent, hydrofluoric acid is repeatedly filtered
through quartz, and lead carbonate is added to the result-
ing fluosilicic acid, until the solution contains 60-90 gms.
of lead per 100 gms. of free fluosilicic acid. The optimum
temperature lies between 30 C. and 35 C.

Pring 60 gives the following suitable electrolyte : —

H2SiF 6 . . . . 9*5-10*5 per cent.

Pb as PbSiF 6 . . . . 4'5-5'2 per cent.
Sp.gr. .. .. 113-1-16

It has been found necessary to add a small quantity of
colloid such as glue or gelatine not exceeding 0'i per cent.

88 INDUSTRIAL ELECTROMETALLURGY

and generally about 1/5000. Owing to its destruction at
the anode, where it prevents the formation of lead peroxide,
frequent small additions are necessary. At Trail, B.C.,
0*007 P er cent, of glue is added every other day. The
following analyses are typical of the deposited lead and the
slimes : —

Impurities in the lead. Slime.

Cu 0*0010 per cent. Pb 10*3 per cent.

Bi 0*0022 „ Ag 47

As 0*0025 „ Sb 25*32

Sb 0*0017 „ As 44*58

Bi 0*52

Cu 9*3

Betts has suggested the use of other addition agents in
addition to glue and gelatine, such as pyrogallol, phenol,
resorcin, and saligenin, including anodic depolarizers like
sulphurous acid, hydroquinone, and o.amidophenol. The
current yield is said to be from 85 to 90 per cent. Senn 61
and Kern 62 confirmed the utility of Betts' electrolyte.
Fischer, Thiele and Maxted 63 also obtained good deposits
with fluosilicates, fluoborates, fluozincates, and fluostannates.

Various other electrolytes, in addition to fluosilicic acid
salts, have been suggested and are the subject matter of
numerous patents. Siemens and Halske w have patented
the use of lead perchlorate containing free perchloric acid
and an organic colloid as an electrolyte. It is said that the
Hagener accumulator works are using this electrolyte on
a large scale. Peptone appears to be the best addition agent
for perchlorate baths, although mucilage, albumen, salep,
and other vegetable mucilages have been patented by the
same firm.

A suitable bath was found in an electrolyte containing

Pb as perchlorate, 5 per cent.
HCIO4, 2-5 per cent.
Peptone, 0*05 per cent.

A current density of 2-3 amperes per sq. dcm. at a voltage
of 0*21 with electrodes 2*5 cms. apart, gave solid smooth
deposits with a current efficiency of over 99 per cent.

ELECTROLYSIS IN AQUEOUS SOLUTIONS 89

Mathers and Overman w found the most suitable addition
agents in order of merit to be —

Clove oil 100 c.c. per ton of lead deposited.

Peptone 350gms. „

Phloridizin.

Snowden 66 modified the Tommasi process by using a
cathode rotating at high speed and o*i per cent, of gelatine
in the acetate electrolyte. The use of nitrate solutions as
well as complex electrolytes, such as lactates and oxalates,
have been investigated, but the deposits obtained from the
solutions are not as good as those from the electrolytes
already enumerated.

The electroplating of metals with lead as protection
against acid corrosion with the above electrolytes has not
come up to expectation.

Difficulties have been encountered in the satisfactory
treatment of the slimes recovered in the electrolytic lead-
refining plants. The slimes contain lead, arsenic, and
antimony, with smaller traces of copper, iron, silver, and
more rarely bismuth, gold, and tellurium. One of the most
satisfactory methods of dealing with this complex is the one
adopted at Trail. After washing with water and weak
alkali to remove the last traces of acid the slimes are boiled
in a 6 per cent, sodium sulphide solution, containing about
1 per cent. Na^ Antimony is thus removed and recovered
by electrodeposition (see p. 90). The slimes are then
leached with hot sulphuric acid in the presence of air. From
the solution the silver and copper are removed and gold
recovered from the residue. Other methods, such as ex-
traction of the antimony with hydrofluoric acid, to which
is then added sodium potassium fluoride and the antimony
recovered by electrodeposition, whilst the residues are
subjected to chlorination and fractional electrolytic precipi-
tation, amalgamation processes or casting the slime into
anodes with subsequent electrolytic treatment, have all
been suggested, but details of technical working are lacking
for the majority of these suggestions.

go INDUSTRIAL ELECTROMETALLURGY

Antimony.

The electrolytic deposition of antimony has been
developed on a technical scale b} r Siemens and Halske.
As electrolyte a solution of antimony sulphide in sodium
sulphide is used, the antimony sulphide ore being leached
with the spent electrolyte. In the original process a
divided cell was used, the antimony being deposited from
the circulating catholyte on sheet iron cathodes, whilst in
the anode compartment where carbon anodes are placed,
chlorine is liberated from a salt solution. At Trail, where
lead slimes are used as a source of antimony, the divided
cell is dispensed with, and the sodium sulphide is allowed
to be partially oxidized at lead anodes to sodium sulphite
and sulphate. With a current density of 07 amp. per
sq. dcm., with an applied E.M.F. of about 1 volt and an
electrolyte temperature of 6o° C, antimony practically
pure is deposited as a dull warty sheet about 3 mm. thick.
The deposited metal is removed by melting under a flux
of soda and potassium sulphide, which effectually removes
the last traces of sulphur, and cast into ingots showing
the characteristic stellate crystalline structure.

A 6 per cent, solution of sodium sulphide is used as solvent
and electrolyte; antimony pentasulphide dissolves in this
solution as follows : —

SbgSg +3Na 2 S =2Na 3 SbS 4
which partially dissociates into the following : —

NaaSbS^Na+SbS*'"
Since the complex SbS'"* is not readily dissociated again —

SbS /// 4$Sb-+2S'+2S ,/

the equilibrium of Sb"" ions in a 6 per cent, solution of
antimony sodium sulphide being only io~ 60 w (EaSb/N.Sb"*
==-0-463, whilst E A Sb/£Sb- in Na2S=+0709 volt).
Ost and Klapproth 87 assumed that the deposition of antimony
was caused by the secondary reaction caused by the discharge
of sodions at the cathode as follows : —

5Na +Na 2 SbS 4 =Sb +4Na 2 S

ELECTROLYSIS IN AQUEOUS SOLUTIONS 91

Whether the mechanism is a direct electrodeposition of
antimony or is caused by a secondary decomposition, there
is always an anodic liberation of free sulphur. Free sulphur
reacts with sodium sulphide to form the polysulphide—

NajS+S^NaaSg
which on diffusion to the cathode will dissolve antimony
to form a thioantimonate —

2Sb +3Na2S 2 =2Na 2 SbS3

Consequently only a low current efficiency can be claimed
in a single cell process such as is used at Trail, unless a
sulphur depolarizer is added to the electrolyte, the average
efficiency lying between 45 and 50 per cent. Among the
more important sulphur depolarizers which have not yet re-
ceived technical application may be mentioned sodium
sulphite and potassium cyanide —

Na2S 2 +Na2S0 3 =Na2S 2 8 +Na 2 S

NagSa +KCN =KCNS +Na 2 S

Experiments have also been conducted with other electro-
lytes in addition to the alkaline sulphides. At Newcastle-
on-Tyne antimony is deposited from a solution of the
fluoride in an electrolyte of hydrofluoric acid containing
potassium and sodium fluoride. Betts 68 has suggested the
use of acid solutions containing iron salts in a divided cell,
the ferric salts generated #nodically being used to dissolve
more antimony from the slimes. Successful electrolytes
were found in the mixtures of antimony trichloride and
trifluoride w ith the addition of ferrous sulphate or chloride.

Bismuth.

The technical electrodeposition of bismuth has not been
successfully developed on a large scale. Although the
electrolytic potential of bismuth in a solution containing
its ions lies considerably below the point where hydrogen
evolution should commence Ea= — 0*393 volts, yet, owing
to the tendency for this element to form complex salts in
solution hydrogen, evolution is unavoidable. Under these
conditions bismuth is deposited, either in a spongy condition

92 INDUSTRIAL ELECTROMETALLURGY

or as closely adherent crystals, with a very low current
efficiency. Foerster and Schwabe 6g claim to have obtained
good deposits from a fluosilicate solution, whilst Sand 70
obtained deposits suitable for electroanalytical work from
nitrate solutions.

Excellent deposits may be obtained from sodium tartarate
and oxalate solutions provided that the cathode potential
is carefully adjusted continuously during the deposition.

The regulation of the cathode potential could possibfy
be eliminated if a divided cell were used, in which a constant
anodic depolarisation under a constant current density and
a carefully regulated terminal voltage could be maintained.

Tin.

There has been no electrolytic process devised for
the winning and refining of tin. The usual metallurgical
methods are sufficiently simple and economical (m.p, Sn
=233° C), and the impurities in crude tin, chiefly lead,
antimony, and iron, with but small traces of silver and gold,
are not sufficiently valuable, totalling only i to i'5 per cent.,
to warrant an electrolytic refining process. Various ex-
traction processes have been the subject of patent literature,
but have not become technically successful, amongst which
may be mentioned —

Fusion Processes. — (A) Fusion of the ore with caustic
soda and subsequent leaching with water and electro-
deposition, according to Goldschmidt's process.

(B) Fusion with soda and sulphur and subsequent
leaching with water and electrodeposition from the thio-
stannate solution according to Claus's process.

Leaching Processes. — (A) Alkaline leaching with
caustic soda or caustic soda containing sodium sulphide.

(B) Acid leaching with sulphuric, hydrochloric of acid
ferric chloride solutions.

The recovery of tin from scrap iron plate has, however,
become an important electrochemical industry, and has
led to an investigation into the most suitable conditions
for the deposition of tin. Although the Goldschmidt

ELECTROLYSIS IN AQUEOUS SOLUTIONS 93

chlorine stripping process is extensively employed and with
the growing supply of chlorine gas at low prices is likely
to extend, j r et the electrolytic processes have been developed
and are as yet holding their own. Before the war over
30,000 tons of tin scrap per annum found their way to
Germany for detinning. Tin plate averages some 2*5 to 5
per cent, tin by weight, and the residual iron is in great
demand for electric furnace steel work. The more important
electrolytic detinning processes may be classified as follows : —

A. Alkaline Electrolytes.

(1) Beat son's Process 7 1 developed by Goldschmidt. 72 The
scrap tin plate is compressed, perforated, and washed with
caustic soda to remove fats and paint. About 15 kilos of the
clean tin scrap is loosely packed in an iron cage and suspended
in an iron tank which serves as a cathode. The electrolyte
is an 8 per cent, caustic soda solution, and must be regene-
rated from time to time, since it is continually being used
up by absorption of carbon dioxide ; when the concentration
of alkali becomes too low stannic hydroxide separates from
the electrolyte.

The temperature of the electrolyte is maintained at
70 C. by steam heat, and the current density of o*8o to
1 amp. per 100 sq. cms., with an E.M.F. of 17 volts (which
rises to 2*5 volts when detinning is complete). The tin
is deposited from the solution in a spongy form containing
a little copper, iron, and lead with an 80 per cent, current
efiiciency (assuming solution and deposition of tetravalent
tin). The sponge is compressed and melted with coke.
Foerster and Dolch investigated the mechanism by which
the tin is dissolved at the anode and deposited on the
cathode. 73

Dissolution and precipitation of the tin in the tetra-
valent state have been shown to take place —

Sn->Sn""

with a current efficiency of 80 per cent., but it appears more
probable that dissolution takes place as follows : —

Sn->Sn"

94 INDUSTRIAL ELECTROMETALLURGY

the divalent alkaline stannite being anodically oxidized by
the oxygen liberated. Tin becomes readily passive in
alkaline solution owing to the formation of a film of colloidal
stannic hydroxide; when this occurs the anodic potential
is raised sufficiently to cause the evolution of oxygen.
Cathodic reduction of Sn" ,# to Sn" before deposition does
not appear to take place.

Gelstharpe 74 favours agitation of the electrolyte, which
reduces the applied E.M.F. for stripping and deposition
by about 0*5 volt. Borchers 76 and Keith 76 advocated the
addition of sodium chloride to the alkaline electrolyte. If
more than 3 per cent, of sodium chloride is added iron is also
dissolved. Sodium nitrate as well as sodium cyanide have
been advocated as addition agents with unsatisfactory results.

(2) Borchers' Process. — Borchers proposed an electrolyte
containing 15 per cent, sodium chloride and 3 per cent, of
sodium stannate as electrolyte. With a temperature of
50 C. and a P.D. of 2-3 volts per cell, tin could be effectually
stripped and deposited with a current density of 0*5 to 1*5
amps, per 100 sq. cms. Luckow advocated a fluoride bath
for the same purpose.

(3) Claus's Process. — As electrolyte, a solution of sodium
thiostannate was used, 77 containing 4-5 per cent, tin and
10 per cent, of free caustic soda. Electrolysis takes place
in a warm electrolyte at 8o° C. with sheet iron cathodes and
a current density of 3-4 amperes per 100 sq. cms. All
impurities except arsenic and antimony are removed as
slimes. Steiner 78 advocated the addition of 1 per cent, of
flowers of sulphur to the electrolyte.

B. Acid Electrolytes.

(1) The Neil and Brown Process. 19 — The scrap tin plate
is stripped in boiling ferric chloride solution according to the
equation —

2FeCl 3 +Sn->2FeCl 2 +SnCl 2

The disadvantage of this process is the simultaneous solu-
tion of iron during the period of immersion. The electro-
lyte is circulated through divided cells of concrete first
through the cathode, then back through the anode com-

ELECTROLYSIS IN AQUEOUS SOLUTIONS 95

partments. The cathodes are sheet tin, and separated from
the graphite anodes by earthenware diaphragms. At the
cathodes tin is deposited —

Sn"->Sn+2®

whilst at the anodes the ferric chloride is regenerated —

2FeCl 2 4-2C1' =2FeCl 3 +20

Provided that the tin plate could be stripped without

simultaneous solution of iron this process would be more

economical than the Goldschmidt one. Hemingmay 80 uses

ferric sulphate as a leach. Divided cells are not used, but the

ferrous sulphate is reoxidized by sodium nitrate.

(2) The Bergsoe Process. — Cold tin tetrachloride is

used as stripping solution, the tin going into solution as

follows : —

SnCl 4 +Sn->2SnCl 2

Tin cathodes and graphite anodes are used. The process
is open to the same objection as the Brown, namely the
simultaneous solution of iron with the tin. Rienders 81
conducted experiments on stannous chloride and stannic
acid solutions as electrolyte with the addition of ammonium
chloride. Solution of the tin proceeds both chemically and
electrochemically in the stripping cell, and the excess of
tin in solution is subsequently removed in separate cells,
utilizing graphite anodes. A current density of 1 to 2
amperes per 100 sq. cms. is employed.

Gelstharpe 82 carried out successful experiments at
Manchester with a 1*25 per cent, solution of hydrochloric
acid containing a trace of sulphuric acid as electrolyte ;
with a current density of 17 amperes per 100 sq. cms. at
i'5 volts practically pure tin sponge could be obtained.

Sulphuric Acid stripping and depositing electrolytes
have been suggested by Smith and Englehardt, 83 the latter
claiming a current efficiency of over 60 per cent. Nauhardt
suggested the addition of a small quantity of ammonium
sulphate. A good deposit was obtained with a current density
of 02 to 03 ampere per sq. dcm. Quintaine 84 deposited tin

96 INDUSTRIAL ELECTROMETALLURGY

from a sulphate solution on lead cathodes. Nodin 85 used
sulphuric acid as a stripping agent, followed by electro-
deposition in separate cells on the basis of the Neil and Brown
process.

C. Miscellaneous Electrolytes.

Matuschek 86 has suggested the use of ammonium oxalate
dissolved in a saturated solution of tin ammonium chloride
as a suitable stripping and depositing electrolyte. Good
deposits could be obtained at current densities as high as
3 amperes per ioo sq. cms., provided some colloidal addition
agent were employed. Tannin, gum, and a small quantity
of NaH 2 P0 4 were stated to be most suitable. Hollis 87
suggested the use of tin fluosilicate as a suitable electrolyte.
Mennicke 88 observed that the best conditions for deposition
were obtained with an electrolyte containing 10 per cent,
tin and 10 per cent, free hydrochloric acid. Electrolysis
was conducted with a current density of i ampere per
ioo sq. cms. at 20° C.

The alkaline electrolytes suffer in practice from their
instability in presence of atmospheric carbon dioxide, and
the fact that the iron tin alloy formed at the junction of
the two metals is not dissolved. The whole of the tin is
removed by acid electrolytes, but the simultaneous solution
of the iron which has already been referred to renders
these stripping agents even more unsuitable than the alkaline
ones.

Tin Plating. — There are many difficulties associated
with the electrodeposition of tin as a white dense adherent
deposit. Not only do the anodes dissolve irregularly in
excess of the amount deposited on the cathode, but the
deposited metal is generally dull, powdery, and loosely
adherent. Special precautions as regards cleanliness of
the surface which is to receive the deposit have to be taken.

Iron and steel are generally given a thin copper deposit
before the tin coat to ensure adherence of the tin, due to
the formation of alloys, CugSn, Cu^n. Very low current
densities must be employed, and as electrolytes those which
form complex ions are found most suitable. For good

ELECTROLYSIS IN AQUEOUS SOLUTIONS 97

deposition high temperatures and efficient circulation of
the electrolyte are essential. Thick, dense deposits can only
be obtained by rotating the cathode at high speed or by
removing the electrode from time to time and scraping the
deposit with a fine wire brush.

Alkaline Electrolytes. — Twenty-five grammes of
stannous chloride dissolved in a litre of water containing
60 gms. of caustic soda or 20 gms. of caustic potash forms
a suitable electrolyte. With a current density of O'l ampere
per 100 sq. cm. good deposits may be obtained. Steel and
Eisner 89 recommended the addition of potassium cyanide
to the electrolyte. In the Brass World 90 the following
electrolytes for giving good deposits on brass and iron are
stated : —

(1) On Brass —

Gms. per litre.

KOH 7-5

SnCl2 . . • • • • • • 75

KCN 350

(2) On Brass or Steel —

KOH • • • • • • • • 15

SnCi2 • • • • • • • • 15

KCN 35

It is recommended to use the solutions warm and electrolyze
with a bath voltage of 2' 5 to 3 volts. A large anode surface
is desirable.

Acid Electrolytes. — The use of acid oxalates and pyro-
phosphates in acid solution form the basis of a great number
of electrolytes for tin deposition. Roseleur's electrolyte
is the most generally used, and gives satisfactory deposits.
Pure tin anodes must be employed, and the electrolyte
containing 125 gms. of sodium pyrophosphate and 1*5 gms.
of stannous chloride per litre must be kept hot. Field 9l
mentions an oxalate bath of the following composition : —

Grammes per litre.

Stannous chloride . . . . 25-30
Acid ammonium oxalate . . 55-65
Acetic acid . . . . . . 3-4

*<• 7

98 INDUSTRIAL ELECTROMETALLURGY

The bath is conveniently worked at 65 C. with a current
density of 1 ampere per 100 sq. cm. Other solutions con-
taining tartaric and lactic acids have also been suggested.
Kern 92 gives a r&ume of the work published on the deposition
of tin and has further investigated the effect of addition
agents in the nature of the deposit. Tannin in the propor-
tion of 1 gramme to 1*5 litres of solution was found to be the
most beneficial in stannous chloride and fluoride solutions.

Nickei,.

The electrolytic recovery of nickel from its ores, chiefly
sulphide and arsenide, is associated with difficulties, inas-
much as the nickel ore always* contains relatively large
quantities of copper and iron. Attempts to use nickel
matte anodes in a nickel sulphate or chloride electro-
lyte have not proved technically successful, although
Giinther 93 obtained good and uniform solution of such
electrodes in a sulphate solution. The sulphur is liberated
in a free state at the anode. W. Trumm 9 * developed a
process for the Orford Copper Co. using nickel sulphide
electrolytes in a nickel chloride solution. It is said that the
process proved satisfactory on a small scale.

As in the case of copper either desulphurization of the
matte, or extraction processes are necessary to avoid unduly
fouling the electrolyte. The Canadian Copper Co. have
experimented successfully on a desulphurized nickel matte
containing both copper and iron, casting the same into
anodes. As electrolyte, a chloride solution was used,
obtained by chlorine treatment of desulphurized matte in a
brine solution. Electrolysis was conducted in a series of
concrete vats ; in the first series, with an applied E.M.F. of
0*35 volt, copper was deposited on electrolytic copper
sheet cathodes. When the ratio nickel to copper in the
electrolyte exceeded 80 : 1 the rest of the copper was pre-
cipitated by sodium sulphide, the iron removed as hydroxide,
and the bulk of the salt removed by concentration. The
nickel was finally removed by deposition on nickel sheet

ELECTROLYSIS IN AQUEOUS SOLUTIONS 99

cathodes, using graphite anodes enclosed in earthenware
diaphragms to remove the chlorine. With an applied
E.M.F. of 3*5 to 3*6 volts, nickel over 99*85 per cent, in
purity could be deposited with a current efficiency of 93
per cent.

Extraction Processes. — Hoepfner (see p. 35) modi-
fied his electrolytic process for copper ores which has already
been discussed, to nickel. After roasting the ore to render
the iron insoluble, extraction of the copper and nickel
sulphides was accomplished by means of a solution of
cupric chloride containing calcium chloride according to
the equation —

NiS +2CuCl 2 ->Cu 2 Cl 2 +NiCl 2 +S

The silver and iron having been removed chemically and
the copper electrically, the electrolyte containing but little
copper and all the nickel was passed on to cells of similar
construction as used for removing the copper, but a nickel
sheet cathode was substituted for a copper one. The
graphite anode was depolarized by the returning cuprous
and nickelous chloride solutions.

The separation of copper and nickel can be made nearly
complete by adjustment of the cathode potential, as is
indicated by the following figures for the cathodic potential
equilibrium values between the metals and their solutions : —

Ni/tt Nisalt E A = +0228 volt
Cu/w Cii salt E*=— 0308
Fe/w Fe salt E* =+0*340

This process was modified by Wannschaft 96 in that the
roasted ore was treated with chlorine after being ground
with a calcium chloride solution, a further quantity of
ground ore being added when the solution is heated to 6o° to
70 C. The iron in solution is removed as ferric hydroxide
by agitation with air, and the liquid after filtration contains
about 100 gms. of nickel per litre as NiCl 2 . Nickel sheet
cathodes and carbon anodes are employed with a current
density of 1-1*2 amps, per 100 sq. cm., and 4-4*5 volts per
cell, a current efficiency of 93 per cent, is? seated to have been

ioo INDUSTRIAL ELECTROMETALLURGY

obtained. The chlorine liberated at the anodes was col-
lected by means of hoods. Analyses of the deposited nickel
showed only traces of impurities, o*o6 per cent. Fe, 0*02
per cent. Cu, and 0*02 per cent. Si0 2 . It is stated that crude
nickel copper alloys obtained by desulphurization of the
sulphides can be successfully refined in an acid copper
sulphate electrolyte maintained at 30 C. After the copper
is removed nickel can be recovered by electrolysis at a higher
applied E.M.F. with insoluble anodes. Details of these
processes are, however, lacking.

The Electrolytic Refining and Plating of Nickel.
— It has already been indicated that practically complete
separation of nickel from copper can be obtained by careful
adjustment of the cathode potential. The electrolytes
favourable for the deposition of copper are, however, not
those from which nickel can be deposited successfully.

Since the cathodic potential of nickel in a normal nickelic
salt solution is +0*228 volt, it follows that hydrogen would
be more easily liberated than nickel out of even a moderately
acid solution. The difficulty is further emphasized by the
fact that the overpotential of hydrogen on nickel is low
according to Caspari, less than 0*20 volt, and that the
velocity of reaction —

Ni"«*Ni+2©

is very slow. 96 Nickel and iron have a marked tendency to
become anodically and cathodically passive, thus necessi-
tating an increased cathodic polarization. With a working
current density of ro amperes per 100 sq. cm. a cathode
potential difference of — 0*64 volt was found necessary.
It follows that a nearly neutral solution for the electrolyte
is most desirable, provided that the formation of basic salts
is avoided. In common with other metals that easily become
passive, such as gold in a chloride solution and iron, the
velocity of solution of the nickel anode and of deposition
of metallic nickel from the ionic condition are greatly
accelerated by rise in temperature. 97 Accordingly the best
conditions for deposition are found at relatively high

ELECTROLYSIS IN AQUEOUS SOLUTIONS 101

temperatures, viz. 6o°~70° C, at a high concentration of
nickel ions, and a solution as nearly alkaline as can be con-
veniently managed without the deposition of basic salts.

Nickel Plating. — The advantages to be obtained by
a fine deposit of adherent and dense nickel on metals are
partly negatived by the difficulties inherent in the methods
of electrodeposition employed. The inclusion of relatively
large quantities of hydrogen and probably small quantities
of iron 98 cause the deposited nickel to become brittle and
hard and exhibit a great tendency to peel. Better deposits
may be obtained at high temperatures.

Nickel does not give a satisfactory deposit on zinc or
tin unless a " Striking " bath is employed, more commonly
a thin deposit of copper is first formed before the nickel is
plated on. Cast nickel anodes are preferable to rolled or
electrolytic nickel in the usual electrolytic deposition baths,
since they exhibit only a small tendency to exhibit passivity
phenomena ; this may be counteracted by the addition of
small quantity of nickel chloride to the bath or by the use
of chloride electrolytes. When thick deposits are required
the nickel plating bath must be run warm about 70 C,
but for ordinary thin deposits room temperature is usually
maintained. Of the various electrolytes suggested for the
deposition of the nickel the following have been shown to
be most successful.

Sulphate Electrolytes. — Brochet modified Pfanhauser's
solution " for the composition of a nickel ammonium electro-
lyte-
Nickel sulphate, 166 gms. per litre.
Nickel ammonium sulphate, 55 gms. per litre.

The electrolyte is conveniently operated at room tempe-
rature with a current density of 03 ampere per 100 sq. cm.
The alkalinity of the bath decreases when relatively in-
soluble anodes are employed, and must be corrected. A
hard good deposit is obtained suitable for iron or steel.
A softer and thicker deposit may be obtained by substi-
tuting ammonium citrate or tartarate for the nickel sulphate
in the above electrolyte.

loz INDUSTRIAL ELECTROMETALLURGY

Chloride Electrolytes. — Nickel chloride (15 gms. per
litre) gives an unsatisfactory deposit unless converted into
the double salt nickel ammonium chloride when deposits
equal to those obtained from the double sulphate electrolyte
may be obtained. Dechert has suggested the use of calcium
chloride as a substitute for the addition of ammonium
chloride.

Other acid complex electrolytes have been used from time
to time. Pott's electrolyte containing nickel acetate
(20 gms. per litre), calcium acetate (16 gms. per litre) and
glacial acetic acid (3 gms. per litre), is stated to be an ex-
cellent electrolyte for the deposition of the metal.

Pfanhauser and I^angbein both recommend the addition
of boric or citric acid to the double sulphate electrolyte,
whilst Powell 10 ° suggested benzoic acid.

Nickel ethyl sulphate, 101 nickel phosphate with sodium
pyrophosphate, 102 nickel fluosilicate with aluminium fluo-
silicate, and ammonium fluoride 108 are found among the more
recent patents in various dilutions as suitable electrolytes
for the deposition of dense and smooth deposits on zinc or
brass.

It is claimed that malleable nickel may be deposited
from either of the following electrolytes 104 : —

(1) 8 per cent, nickel as nickel fluoborate.

(2) NiCl 2 5 per cent. ; nickel borate 2 per cent.

It will be noted that only very weak acids are suitable
as addition agents and that the best deposits are obtained
from very nearly neutral electrolytes.

Alkaline Electrolytes. — A few suggested electrolytes
contain nickel as the complex ion Ni(NH 3 )" 4 , amongst which
may be mentioned —

1. Desmur's solution —

Nickel ammonium sulphate 7 gms. per litre.
Sodium bicarbonate . . 8

2. Bischof's solution —

Nickel sulphate . . . . 86
Ammonium sulphate . . 17
Ammonia (o'88o) . . . . 120

n ft

tt t>

ELECTROLYSIS IN AQUEOUS SOLUTIONS 103

The complex cyanide solutions have proved unsatis-
factory for nickel deposition. Certain organic addition
agents have been recommended for ensuring smooth even
deposits. Tannin, gelatine, glue, certain glucosides and
glycerine have all been the subject of patent literature,

Cobai/t.

The electrolytic preparation or refining of cobalt from
its ores has not been the subject of technical investigations.
Doubtless, methods applicable to the deposition of nickel
could be adapted to suit this metal on account of their close
similarity; the electrolytic potential of cobalt E*=* +0*232
being only +0*004 vo ^ higher than that of nickel. Owing
to the lack of demand for this element the price rules higher
than that for nickel, although the available supplies are
large.

Recently, experiments on electroplating with cobalt
have indicated that this metal apparently ofiers some
advantages over nickel deposits. O. P. Watts 105 has sum-
marized the somewhat conflicting evidence in respect to
the merits of the two metals. Kalmus, Harper and Savell, 106
as a result of a long series of technical experiments, came to
the conclusion that cobalt plating was superior to nickel
for the following reasons : —

(1) Cobalt ammonium sulphate is 2*5 times as soluble
as nickel ammonium sulphate, thus permitting of a greater
speed of electroplating with the same applied E.M.I?.

(2) The cobalt film was strongly adherent and hard
on both brass and iron.

(3) A current up to 4 amperes per sq. dm. can be em-
ployed continuously in cobalt plating baths which is over
three times the current density permissible with nickel.
In one electrolyte a current density of 26*4 amperes per sq.
dm. was used for a short period, and produced a satisfactory
deposit.

(4) The deposited cobalt is harder than nickel, it takes
a high polish showing a beautiful white lustre with a slightly

104 INDUSTRIAL ELECTROMETALLURGY

bluish tint. The actual weight of hard metallic cobalt is
computed to give the same protection as 4 times its weight of
the softer nickel.

(5) Both cast and rolled cobalt anodes may be used ;
passivity phenomena do not appear to be so much in evidence
in the electrolytes employed by these investigators.

(6) Plates up to any desirable thickness may be de-
posited.

(7) Current efficiencies of nearly 100 per cent, could be
obtained with current densities up to and over 5 amperes
per sq. dm.

The two most satisfactory electrolytes were found in
baths of the following compositions : —

(1) Cobalt ammonium sulphate (cryst), 200 gms. per
litre.

(2) Cobalt sulphate, 312 gms. per litre.
Sodium chloride, 19*6 „ ,,
Boric acid, nearly to saturation.

Cobalt is also probably superior to nickel owing to the
fact that hydrogen is much less soluble in the former metal,
and we have noted that the peeling properties of metal films
can generally be attributed to the solution of this gas in the
metal.

The cobalt ammonia electrolytes containing the complex
ion Co(NH) 3 )' # 4, suggested by Boettger, Beardslee, and
others, have not proved satisfactory in practice.

The double sulphate bath mentioned above has been
modified by the addition of magnesium sulphate with or
without a small quantity of citric acid.

In practice the use of baths weaker than (1) and (2)
would be indicated owing to the unavoidable loss of solution
on removing the plating articles. Langbein suggests as
a depositing bath suitable for electrolysis —

Cobalt ammonium sulphate 40 gms. per litre.
Boric acid . . . . 20 ,, ,,

Deposition of Cobalt Nickel Alloys. — O. P. Watts
gives the composition of a bath from which it is claimed

>> )9

»l 99

ELECTROLYSIS IN AQUEOUS SOLUTIONS 105

the hardest alloy of nickel and cobalt can be deposited
(75 per cent. Ni : 25 per cent. Co) —

Nickel ammonium sulphate 147 gms. per litre.
Cobalt ammonium sulphate 40
Ammonium sulphate . . 56

I^angbein suggests the addition of boric acid. 107

Deposition of the two metals from such solutions in
the ratio of 3 Ni to 1 Co can undoubtedly be obtained,
anodic solution of the two metals must, however, be in the
corresponding ratio. There are two alternative schemes
by which this could be accomplished, either by the inser-
tion of two electrodes, one nickel and the other cobalt, and
passing the correct current for dissolution through each
electrode, or by the casting of an alloyed anode. The
nickel cobalt anode would probably dissolve with perfect
uniformity, since the metals are miscible in all proportions
in solid solution.

REFERENCES TO SECTION II.

1 Metall., 5. 27 ; 1908.

1 Trans. Amer. Electrochem. Chetn. Soc, 25, p. 193 ; 1914.

8 Metall., 8, 820 ; 1906.

4 Trans. Atner. Electrochem. Soc, 190, xxxvii.

6 Metallurgie, 1908, p. 202.

• Blount, " Practical Electrochemistry," p. 44.

7 Addicks, Trans Atner. Electrochem. Soc, 26, 1914.

8 J.C.S.I., 347, 91, 1907 ; Phil. Mag. (6), 1901, 45-79.

• Zeit. Phys. Chem., 28, 1897, 689.

10 Zeit. Phys. Chem., 47, 1904, 52.

11 " Elektroanalytische Schnell Methoden." Stuttgart, 1908.
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18 Trans. Amer. Electrochem. Soc, 21, 236; 1912.

14 Luckow, Zeit. Anorg. Chem., 19, 11 ; 1880.

15 Firchland, Eng. Pat, 24806 of 1906.

16 " The Corrosion of Iron and Steel." E. Rideal. 1913.

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80 /. Phys. Chem., 1916, 20, 296.

81 Trans. Amer. Electrochem. Soc, 27, pp. 291 ; 191 4.
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88 Bunger, Electrochem. Ind., 8, 17; 1905.

84 Chem. Tech., 2, 23.

86 " A Treatise on Electrometallurgy/' p. 258. Macmillan.

86 Chem. Zeit., 28, 1209; 1904,

io6 INDUSTRIAL ELECTROMETALLURGY

87 Met. Ind., 9, 509; 1911.

88 Michailenko, /. Russ. Phys. Chem. Soc, 44, 507; 191 2.
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80 Frary, Trans. Amer. Electrochem. Soc, 1913, xxiii.

81 Langbein, p. 383.

88 Zeit. Anorg. Chem., 55, 321.
88 Zeit. Anorg. Chem., 25, 995 ; 1912.
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86 Zeit. Elektrochem., 11, 575; 1906.

87 Zeit. Elektrochem., 16, 25 ; 1910.

88 Foerster, Eng. Fat. 6276, 1909.

88 /. Amer. Chem. Soc, 28, 1350; 1996.

40 Electrochem. and Metallurgist, 8, 490; 1904.

41 Ber., 1892, 25, 779.

48 Zeit. f. Angew. Chem., 827 ; 1900.
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46 Kern, Met. Chem. Eng., 9, 443.

46 Bancroft, Trans. Amer. Electrochem. Soc, 1904, 6, 200; and Leduc,

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47 Schuster, Proc Phys. Soc, 50, 344.

48 J.C.S., 1906, 22, 43.

48 Trans. Amer. Electrochem. Soc, 1905, 7, 143.

50 Zeit. Elektrochem., 4, 451 ; 1878.

61 /. Electrochem., 9, 979; 1903.

88 J.S.C.I., 23, 754 ; 1909.

88 Dingier, Poly. J., 88, 30; 1843.

64 Ber., 1882, 1, 276.

55 Zeit. Elektrochem., 13, 219; 1907.

58 Zeit. Angew. Chem., 24, 1000; 1908.

87 Electrochem. Ind., 2, 355 ; 1904.

88 See " Lead Refining by Electrolysis," A. W. Betts. 1908.
68 D.R.P., 38, 193 of 1886.

80 " Some Electrochemical Centres." 1908.

81 Zeit. Elektrochem., 11, 229; 1905.

68 Trans. Amer. Electrochem, Soc, 13, 441 ; 1909.
68 Zeit. Anorg. Chem., 67, 302 ; 1910.
64 D.R.P., 223, 668 of 1908.

68 Trans. Amer. Electrochem. Soc, 21, 313; 1912.
66 /. Phys. Chem., 10, 500 ; 1900.

87 Zeit. Anorg. Chem., 1900, 827.

88 Trans. Amer. Electrochem. Soc. 8, 187; 1905.

69 Zeit. Elektrochem., 16, 279; 1910.

70 J.C.S., 1907, 373, 91.

71 B.P. 11,067 of 1885.

78 Met. & Chem. Engineering, 10, 202 ; 1912.
78 Zeit. Elektrochem., 16, 599 ; 1910.
74 Trans. Farad. Soc, 1, 11 1 ; 1905.
78 Zeit. Elektrochem., 7, 34; 1912.

ELECTROLYSIS IN AQUEOUS SOLUTIONS loy

78 Electrochem. 6* Metall. Ind., 7, 79; 1905.

77 Electrical Review, June, 1907.

78 Eng. Pat. 10,230.
78 U.S. Pat. 707,675.

80 Eng. Pat. 8,759.

81 D.R.P., 245, 628 (Metallurgie, 9, 402; 1912).

82 Electro chemist, 1, 278; 1901.
88 Zeit, Elektrochetn., 7, 34; 191 2.

84 Electrochemical Industry, 2, 237; 1907.
86 Eng. Pat. 7,706 of 190.

86 Ger. Pat. 244,567.

87 U.S. Pat. 916,155 of 1904.

88 Zeit. Elektrochetn., 12, 112; 1905.

88 Watt & Phillip, " Electroplating," 191 1, p. 345. Langbein-Brannt,

,c Electrodeposition of Metals," 1907, p. 440.
•• 7. 121; 1911.

81 " Principles of Electrodeposition." 1911, p. 213.
91 Trans. Atner. Electrochem. Soc, xxiii, 191 3.
88 Metall. u. Erz. t 1, 77; 1904.
84 Kershaw, Electrometallurgy ; 1908, p. 236.
86 Billiter, " Die Elektrochemie Wasseriger Losungen," vol. 1, p. 282 ;

1909.
88 Schock and Schineitzer, Zeit. Elektrochem., 15, 602 ; 1909.

97 Foerster, Zeit. Elektrochem., 4, 160; 1897.

98 Engermann, Rev. d' Electrochem. et d'Electrometal. July, 191 2.

99 Electroplating. 1900.

100 U.S. Pat., 229219.

101 D.R.P., 117054.

188 Langbein, " Electrodeposition of Metals," 1909, p. 319.
108 Kern, U.S. Pat., 942719, 1909.

104 Kern, Trans. Amer. Electrochem. Soc, 15, 464; 1909.

105 Trans. Amer. Electrochem. Soc, vol. xxiii. 191 3.
188 Bureau of Mines, Canadian Dept. of Mines, 1915.
107 The Brass World, 1909, p. 208.

BIBLIOGRAPHY TO SECTION II.

"Handbuch der Elektrolytischen Metallniederschlage," G .Langbein.

Leipzig. 5th edit. 1903.
" Monographien fiber angew. Elektrochemie," W. Pfanhauser. "Elek-
troplattierung Galvanoplastik u. Metallpolieriung," W. Pfanhauser.
Vienna, 191 o.
Elektrometallurgie," Borchers. 1896.
Principles of Applied Electrochemistry," Allmand. 1912.
" Practical Electrochemistry," B. Blount. Macmillan.
'• Electrometallurgy," Kershaw. 1908.
" Electroplating," P. Hasluck. McKay, Pa. U.S.A. 1905.
Electroplating," Barclay and Hainsworth.

Electroplating and Refining of Metals," Watt and Phillipp. Crosby
Lock wood. 1902.

108 INDUSTRIAL ELECTROMETALLURGY

" Electroanalysis," E. Smith. Blakeston, Pa.

" Analyse des Metaux par Electrolyse/' A. Holland and L. Bertiaux.

Dumond, Paris.
" Elektroanalytische Schnell Methoden," A. Fischer. Enke, Stuttgart.
Quantitative Analyse durch Elektrolyse," A. Classen. Springer,

Berlin.
Electrolytic Methods of Analysis/' Neumann- Kershaw,
Practical Electroplating/' Bedell.
" Some Electrochemical Centres," G. N. Pring. Univ. Press, Man-
chester.

Journals,

Transactions of the Electrochemical Society.

Transactions of the Faraday Society.

Electrochemist and Metallurgist.

Electrotnetallurgie.

L' Industrie Electrochemique.

L'Electriden.

Metal Industry.

Metallurgical and Chemical Engineering.

Engineering and Mining Journal.

Zeitschrift fur Elektrochemie.

Elektrochemische Zeitschrift.

Electrical Review.

Electrochemical Industry.

Mineral Industry.

Special Literature.

" Modern Electrolytic Copper Refining," Ulke. Wiley & Sons. 1903.

" Metallurgy of Tin," Louis. 191 1.

" £>ie Darstellung des Zinks auf Elektrolytischem Wege," Gunther.

Knapp. Halle.
" Lead Refining by Electrolysis/' A. W. Betts. Wiley & Sons.
' ' Elektroly tische Verzinkung, ' ' S. Cowper Coles. 1 905.
"Die Metallurgie des Zinns/' H. Mennicke.
" Elektrometallurgie des Nickels," W. Borchers.

Section II.— ELECTROLYSIS IN FUSED

ELECTROLYTES

Sodium.

Pracxigai&y all the sodium produced at the present time
is made by electrolysis of fused caustic soda, although

Gas rings Co eommtnot
fusion.

Fig. 7. — Castner cell for electrolysis of fused caustic soda.

attempts to use sodium chloride as electrolyte have been
partially successful.

The Castner Process. — The Castner cell (Fig. 7) con-
sists of a cast-iron vessel, D, into which an iron cathode, A,
is luted .by fused caustic soda being insulated by a porcelain
ring, E. The ring-shaped anode C insulated from the vessel
and enclosing the cathode is of nickel and usually perforated

no INDUSTRIAL ELECTROMETALLURGY

to permit of free circulation of the electrolyte. Above the
cathode is a ring of nickel wire gauze, B, dipping under
the surface of the electrolyte. The sodium liberated at the
cathode floats to the top and is retained by the wire-gauze
screen. The metal can be ladled out by means of a per-
forated spoon, or a discharge pipe is fitted to the hood. 1
The largest cells are about 60 cm. deep and 45 cm. diameter,
holding about 100 kgm. of molten soda. The cathode
current density is about 200 amperes per sq. dcm., and
the anode density 170 amps, per sq. dcm. at 5 volts, the
total current per cell being 1200 amps., giving a current
efficiency of about 45 per cent. The electrolyte is main*
tained fused by the current, and just sufficient lagging is
placed round the cell to ensure the formation of a thin
protecting crust of caustic soda and a good seal for the
cathode : the cell can be started up by means of a gas
burner. Electrolysis is conducted at as low a tempe-
rature as possible, 3i5°-320° C. Above 525 C. the yield
is practically zero (m.p. crude NaOH 300 C), due to
the increased diffusivity of the metal in the electrolyte.

The Mechanism of Electrolysis. — Le Blanc and
Brode 2 investigated the mechanism of electrolysis and
showed that the electrical current efficiency could never
exceed 50 per cent, owing to the simultaneous liberation
of hydrogen at the cathode according to the equations —

2NaOH=»2Na+20H 1

4OHI at the anode->2H 2 0+0 2 +4©

2H 2 0->2H 2 +0 2 on electrolysis

Net reaction 2NaOH=Na 2 +H 2 (cathodic) +0 2 .
See also V. Hevesy, Zeit. Elcktrochem., 15, 539 ; 1909.

Both the liberated sodium at the cathode and the water
formed at the anode difEuse through the bulk of the electro-
lyte and there react, liberating hydrogen; since metallic
sodium diffuses more rapidly than water at high tempe-
ratures, both hydrogen and oxygen may be liberated in
the anode compartment, causing explosions.

Further reactions between the liberated sodium and

ELECTROLYSIS IN FUSED ELECTROLYTES in

oxygen resulting from the electrolysis may also account
for a small efficiency loss, according to the equation —

2Na+0 2 =Na 2 2

the peroxide being then again reduced by the sodium at
the cathode.

It is evident that as long as the water produced by
the electrolysis is not removed from the electrolyte as such,
but decomposed into hydrogen and oxygen, the current
efficiency can never exceed 50 per cent. Various patents
have been taken out *to effect this removal, eg. by using
a diaphragm unattacked by molten caustic soda to prevent
the water returning to the cathode or by passing dry air
through the anode compartment, but they have not re-
ceived technical application.

The decomposition potential of dry fused NaOH,
according to Le Blanc, 8 is 2*2 volts. Technical electrolysis
is conducted with an applied E.M.F. of 5 volts and a current
efficiency of 45 per cent., giving an energy efficiency of 20
per cent.

Hence 1000 k.w. hours are necessary to produce 79 kgm.
of sodium.

The Griesheim Process. — In this process the " contact
electrode " principle general for production of calcium and
strontium, and occasionally used for preparing magnesium,
is employed.

A circular iron ring in a shallow bath containing the
fused caustic soda serves as anode. The cathode consists
of a vertical iron rod which can be lowered by means of
gearing to make contact with the electrolyte in the centre
of the bath. As fast as the sodium is liberated the cathode
is raised and the end of the sodium rod thus formed serves
as cathode. A cathode current density as high as 1000
amps, per sq. dcm. is claimed for the process, giving a 35
per cent, current efficiency. The chief advantage of the
process lies in the fact that the metal is not so much exposed
to the solvent action of the electrolyte as in the Castner
process. Against this must be set the very high voltage

H2 INDUSTRIAL ELECTROMETALLURGY

necessary to operate a contact electrode process with a high
cathode current density.

Modifications of the Castner Electrolyte. — Becker*
suggested the use of a mixture of sodium carbonate
and soda as electrolyte in a modified Castner cell, which
was provided with a sodium collector above the cathode.
The addition of the carbonate to the caustic soda, how-
ever, raises the melting-point of the electrolyte ; with 50
per cent, carbonate a working temperature of 480 C. is
necessary. Under these conditions the yield of sodium
is, as to be expected, very small, and no carbon dioxide is
evolved at the anode. B. P. Scholl suggested the addition
of 50 per cent, sodium sulphide to the fused caustic soda.
The theoretical decomposition potential of 2*2 volts for the
caustic soda being reduced to i*8 volts.

The free sulphur liberated anodically react9 with the
fused caustic to reform sodium sulphide, which is again

electrolyzed.

Na 2 S=2Na+S"
4NaOH +2S =2NaaS +2H 2 +0 2

It will be noted that although there is a reduction in
the decomposition potential required the fundamental
difficulty, viz. the removal of the water, is not accomplished
by this means.

There are two other salts utilized for the production
of metallic sodium which are worked on a technical scale.

The Darling process (worked at Philadelphia, U.S.A.)
is said to employ fused sodium nitrate as electrolyte. The
central cathode, stated to be made of carbon, is surrounded
by two perforated coaxial metal cylinders, whilst the anode
is the cast-iron containing vessel.

With an applied E.M.F. of 15 volts sodium is liberated
at the cathode and is there recovered in the usual manner by
means of a perforated ladle whilst the anode products from
the annular space between the anode and the perforated
cylinders are removed and converted into nitric acid by
condensation. From the details available of this process it
is difficult to find out how the nitric acid is produced by

ELECTROLYSIS IN FUSED ELECTROLYTES 113

direct condensation, as the anode products would consist
entirely of nitrogen dioxide and oxygen :

2NO' 8 ->2N0 2 +0 2 +2e

The production of nitric acid from this gas mixture by
absorption in water would not offer any advantages over
the Castner process for making sodium and the usual sul-
phuric acid nitre process for strong acid. Liquefaction

D.|C

aso*

3o«P

wo"

*»*

(So

■

»• lo lo 4t ib to 70 flO so

loo

mjcw

Fig. 8. — Melting-point curve of mixtures of Soda and Potash.
(G. v. Hevesy, Zeit. Phys. Chem. t 73, 676.)

of the nitrogen dioxide (see Partington, "The Alkali In-
dustry ") would probably be too expensive even with this
concentrated gas. A direct preparation of sodium and
nitric acid vapour might be obtained by the regulated
admission of superheated steam to the anode compartment,
when the following reactions would conceivably take place :—

4N0 3 ' +2H 2 =4HN0 3 +0 2 +40

If this reaction could be made to proceed smoothly the
preparation of sodium and concentrated nitric acid in one

i*.

H4 INDUSTRIAL ELECTROMETALLURGY

operation would prove more economical than the combi-
nation of the Castner and sulphuric acid distillation process.
The use of sodium chloride for the production of metallic
sodium and chlorine has been frequently attempted. The
processes which have arrived at some technical stage in
their development may be grouped into three classes.

(A) Processes using direct electrolysis between solid
electrodes.

(B) Processes using a molten lead diaphragm serving as
intermediary electrode.

(A) Direct Electrolytic Process. — The preparation
of sodium from fused sodium chloride is scarcely feasible
on the lines of the Castner or Greisheim process, owing to
loss of metal by volatilization, since the m.p. of the electro-
lyte (crude sodium chloride) lies well above 780 ° C, whilst
the liberated sodium has a boiling point of 877 C, and at
8oo° C. has already a considerable vapour pressure.

Early experiments by Fischer on a technical scale
indicated the conditions necessary for the production of
sodium at this temperature. A shallow iron bath divided
into two compartments by a vertical partition extending
nearly to the bottom of the bath was used as the con-
taining vessel. A horizontal carbon anode was disposed
in one compartment and a hollow horizontal metal cathode
placed in the other. By maintaining the temperature of
the metal electrode below that of the electrolyte, sodium
could be drawn off through the tubular orifice. Further
investigations showed that an equimolecular mixture of the
chlorides of potassium and sodium was more suitable as an
electrolyte than the higher melting-point sodium chloride.
The resultant sodium contained about 1 per cent, of
potassium.

The Virginia Electrolytic Company's process, based on
the designs of Seward and V. Kiigelgen plant installed at
Basel, is practically the only one in successful operation.

A circular furnace CC is employed, lined with firebrick,
which is protected by the salt crust EE, and contains a

ELECTROLYSIS IN FUSED ELECTROLYTES 115

circular graphitic anode BB, with a hollow iron cathode A.
The cathode at its upper extremity is surrounded by a water-
cooled hood DD. On electrolysis the deposited sodium
floats up under the water-cooled hood arid flows down through
the circular space into the collecting vessel F. A current
higher than 200 amps, per sq. dcm. cannot be conveniently
used without destruction of the graphite anode. The largest
cell constructed on these lines takes about 10,000 amps.

(KLonns outltl

(B) Process using a Molten Lead Intermediary
Electrode.

The Ashoroft Process 6 is the only one of this type
which has been tried on a technical scale. Several unit
cells absorbing 2000 to 3000 amps, each have been built,
and were stated to function in a satisfactory manner;
nevertheless the process is no longer in operation. The
mechanism of the cell will be seen from the adjoining sketch.
Salt is fed into the cast-iron vessel J, which is provided
with an inner lining of magnesia whilst the temperature
of the vessel is maintained at 8oo° C. The cell is provided
with a molten lead cathode in the base and a vertical carbon

n6 INDUSTRIAL ELECTROMETALLURGY

anode F. The molten electrolyte as well as the molten
cathode is given a rotational movement by means of the
wire helix placed between the magnesia lining and the iron
vessel. .The whole current operating the cell is passed first
through the helix before proceeding to the anode ; in this
way a vertical electromagnetic force field is generated in
the vessel, and since the direction of the current in the
electrolyte can be resolved into both a vertical and hori-
zontal component the magnetic field will cut the horizontal

Fig. io. — Cell for Electrolysis of Fused Sodium Chloride with Inter-
mediary Electrode. Ashcroft Process.

current component at right angles, causing a rotational
movement of the electrolyte.

By means of a suitably situated diaphragm the rotating
molten lead is caused to flow through the orifice I into the
second electrolysis cell K, and return through an annular
space surrounding the first tube D back into the vessel
through the second orifice H. The tube D thus acts as a
heat interchanger for the molten lead ; the second electro-
lytic cell containing fused caustic soda as electrolyte is
maintained at 330 C. On passage of the current chlorine
is liberated at the anode F, and the lead sodium alloy
formed in the first cell is circulated into the second cell,
and returned to the first after the sodium has been removed
and deposited on the iron cathode C. The molten sodium

ELECTROLYSIS IN FUSED ELECTROLYTES 117

liberated at C floats up under the hood B, and is drawn off
through the overflow pipe A.

The cathode current density is stated to be 200 amps,
per sq. dcm. The decomposition voltage of sodium chloride
is about 3*o volts, and should thus be the approximate
working voltage of the cell. In practice the whole system
requires a P.D. of 9 volts. Seven fall over the NaQ cell
and two over the NaOH electrolyte. A current efficiency
of 90 per cent, is said to have been obtained, showing an
energy efficiency of 39 per cent. 1000 kw. hours would,
therefore, produce 85*9 kgm. sodium, a slightly higher
yield than obtained by the Castner plant.

Carrier 6 designed a similar cell to the Ashcroft, but took
no precautions to work the soda electrolyte at low tempe-
ratures. Practically no sodium was deposited at 700 C.

Using a mixture of sodium and potassium chloride as
electrolyte in each compartment, it is stated that a fair
efficiency was obtained with a voltage drop of 6-8 volts
per cell and an anode current density of 20 amps, per dcm.

(C) Processes using a Molten Lead Cathode. —
The earliest experiments on the technical preparation of
sodium were made on these lines, viz. the preparation of
a lead sodium alloy and subsequent fractionation to prepare
pure sodium. These processes are now no longer used to
prepare metallic sodium, but in a modified form, such as
the Vautin, Hulin, and Acker, plants have been largely
developed to produce caustic soda by treatment of the alloy
with steam. 7

Potassium.

The preparation of potassium from potassium hydroxide

can be performed in cells similar to those of the Castner

type. Special precautions must, however, be taken to

protect the liberated metal from oxidation by immersion

in oil.

Magnesium.

Magnesium is prepared by the electrolysis of the fused
double salt of magnesium and potassium chloride, carnallite,
KCl.MgCl2.6H2O, obtained from the Stassfurt deposits.

n8 INDUSTRIAL ELECTROMETALLURGY

Pure magnesium chloride melts at 710° C, but tlie double
salt is easily fused far below this temperature. In technical
operation the electrolyte is maintained between 650 C. and
700° C.

Since molten magnesium is specifically lighter than
fused camallite, it floats to the surface, and has there to be
kept separate from the anodically liberated chlorine. This
is accomplished by means of a porcelain hood, as indicated
in the following sectional diagram. The iron or steel pot C
serves as the container protected from the action of the
liberated chlorine and from the molten electrolyte by a

solidified crust I>. The carbon anode A is inserted in a
porcelain cylinder open at the bottom and having vertical
slits in the part immersed in the electrolyte, whilst an iron
cylinder B immersed in the electrolyte serves as the cathode.

A continuous stream of inert gas (nitrogen or carbon
dioxide) is maintained through the upper part of the cell
during electrolysis, to sweep out any chlorine which may
have penetrated to the cathode compartment.

In practice the temperature is maintained by the elec-
trical energy dissipated in heating the electrolyte. Since
the m.p. of magnesium is 633 C. a somewhat narrow range
is all that can be permitted in working, and difficulties fre-
quently occur due to solidification of the metal.

ELECTROLYSIS IN FUSED ELECTROLYTES 119

Although the decomposition voltage of magnesium
chloride is only 325 volts, yet in practice from 5 to 6 volts
are employed to maintain the temperature of the melt.

If too high voltage be employed an alloy of potassium
and magnesium is formed which readily catches fire and
causes small explosions in the cell. Traces of iron in the
carnallite, a common and nearly unavoidable impurity,
lead to inefficient working due to the alternate reduction and
oxidation of the iron salt at cathode and anode. I^ess than
01 per cent, of ferric chloride can reduce the current
efficiency over 20 per cent, by this means.

Occasionally the small globules of molten magnesium
floating to the surface do not coalesce but are again re-
moved into the electrolyte and are carried as a metal fog
to the anode, where they are reoxidized. This phenomenon,
chiefly due to the formation of a thin oxide film, is caused by
using an inert gas containing oxygen in the cell. By the
addition of a little calcium fluoride, as suggested hy Deville,
the oxide film is dissolved and the magnesium will coalesce.
A fairly high cathode current density is usually employed
from 10 up to 15 amps, per dcm., although A. Oettel 8
has successfully operated a small cell with a current density
as high as 40 amps, per sq. dcm. By careful working a very
high current efficiency can be maintained, over 90 per cent.,
and working with a voltage of 5*5 volts per cell the energy
efficiency is nearly 52 per cent. ; 1000 kw. hours will
produce with a 50 per cent, efficiency 70 kgm. of metal.

The Hemelingen Aluminium and Magnesium Works are
said to use 9 as electrolyte a mixture of sodium chloride and
carnallite in molecular proportions. The process is worked
continuously, and the electrolyte is renewed by the frequent
addition of anhydrous magnesium chloride. Both the
temperatures (750°-8oo°) and cathode current density
(27-30 amps, per sq. dcm.) are higher than usually employed.
The current efficiency is stated to be 70 per cent. Attempts
have been made to make the process more continuous in
its action by reversing the position of anode and cathode
in the containing vessel. The iron rod which now serves

120 INDUSTRIAL ELECTROMETALLURGY

as cathode is slowly raised from the solution, and the molten
magnesium adhering to it solidifies in rod-like form pro-
tected by a coat of fused carnallite, the base of which serves
as cathode in the electrolyte. The control of the cathodic
current density is, however, difficult under these conditions,
and a high potassium content in the metal is usually
obtained.

Tucker 10 has attempted the electrolysis below the melting
point of magnesium at 500° C, when the metal is obtained,
in the form of a sponge, which can be removed and melted
together under a flux of calcium chloride and the electrolyte.

Attempts to deposit magnesium from aqueous electro-
lytes have proved unsuccessful on account of the high
electrolytic solution pressure of the metal Eh =+i'55 volts.
The use of organic solvents for the salts has been the
subject of patent literature, but none have proved of practical
utility.

Calcium.

The preparation of metallic calcium from fused calcium
chloride is more difficult than the production of magnesium,
although the form of electrolyzer employed is essentially
the same in construction.

Pure calcium chloride (m.p. 780 C.) is used as electro-
lyte, although Ruff and Plato 11 and Wohler 12 advocated
the use of a lower melting point mixture of calcium chloride
containing 12 per cent, of calcium fluoride (m.p. 66o° C).
Since the melting point of metallic calcium is 8oo° C. it is
possible by maintaining the electrolyte between 780 C. and
8oo° C. to prepare solid calcium directly by electrolysis. This
is accomplished by means of a contact electrode operated
in the same manner as described above (1).

Borchers and Stockem, 13 who first produced calcium on
a large scale by this method continuously, removed the
calcium and immersed it in petroleum to quench it, whence
a porous residue containing 50 to 60 per cent, metal was
obtained. The metal is fused in a sealed vessel and separated
from the adherent chloride.

ELECTROLYSIS IN FUSED ELECTROLYTES 121

The decomposition potential of calcium chloride is
about 3*25 volts, but in practice very high current densities
must be employed, about 10,000 amps, per dcm., necessitating
an applied E.M.F. of 20-30 volts.

The tendency to metal fog formation observed in the case
of magnesium becomes an important factor in the production
of calcium, and only very small yield9 are obtained unless
the contact electrode process of continuous removal be
employed.

Laboratory experiments on small units have shown,
however, that the preparation of calcium from the fused
chloride can be accomplished with much smaller cathode
current densities than are stated to be used in technical
practice, provided an accurate temperature control is main-
tained. Frary, Bicknell and Tronson H used 9*3 amps, per
sq. dcm. ; Wohler, 15 50 to 250 ; Goodwin, 16 32 to 20 ; and
K. Arndt,* 7 60.

For economical production the temperature in the
neighbourhood of the cathode should just exceed the m.p.
of the metal, but the mass of electrolyte should be as much
as possible below this temperature, but above the point of
fusion of the electrolyte. By maintaining these conditions
the deposited calcium can be made to coalesce round the
cathode, and may be continuously removed in the form of
an irregular rod protected by a layer of fused calcium
chloride without a serious loss as metal fog distributed
through to the electrolyte. The energy efficiency rarely
exceeds 15 per cent. With a 15 per cent, energy efficiency
1000 kw. hours will produce 34*6 kgm. calcium.

Both magnesium and calcium chloride electrolytes suffer
from the disadvantage that in the preparation hydrolysis
may occur resulting in the formation of a hydroxychloride,
which forms an insoluble oxychloride with the liberated
metal. This can be avoided in the initial fusion of the
chloride by the addition of 15 per cent, ammonium chloride
to the moist calcium chloride or carnallite. Regeneration
of an electrolyte containing much oxychloride is stated to
be impracticable. 18

122 INDUSTRIAL ELECTROMETALLURGY

Strontium and Barium.

The manufacture of these elements is only conducted
on a small scale to meet the requirements of chemical
laboratories. The apparatus for their manufacture is,
with some slight modifications, similar to those detailed for
the manufacture of magnesium and calcium. Strontium
and barium do not show such a tendency to produce a fog
as calcium, but appear at the cathode as small molten
drops of metal which coalesce with difficulty.

Lead.

Electrolytic lead refining is usually accomplished in
an aqueous solution (see p. 83), but Borchers has success-
fully refined lead from a fused solution of its salts at a high
current density and electrical efficiency. Although the
direct production of a dense lead without any sponge is a
distinct advantage the method has received no encourage-
ment*

The furnace of cast iron is in two parts, separated from
one another by a water-cooled insulating joint, which is
surrounded and protected by a coating of solidified salt.

The anode side of the electrolytic cell which itself is
placed in the flue of an auxiliary furnace is at an angle,
its inner surface having a series of deep horizontal platforms
which serve to retain some of the crude molten lead fed in
from a hopper at the top. A reservoir in the hearth of the
cell collects the residues of the lead from where it is con-
tinuously run off by means of a syphon. The resultant
lead collects in the hearth on the cathode side, whence it is
removed by a second syphon.

As electrolyte is employed a mixture of lead oxychloride,
potassium and sodium chlorides. The bath is maintained
at about 550 C. With an applied E.M.F. of 0*5 volt and
a current density of loo amps, per sq. dcm. 5 kg. of pure
lead could be obtained per kw. hour. Betts and Valen-
tine 19 obtained a good electrical efficiency, using molten

ELECTROLYSIS IN FUSED ELECTROLYTES 123

lead chloride and sodium chloride as electrolyte, adding
finely crushed galena from time to time. With an applied
E.M.F. of 1 to 1*25 volts good yields of molten lead could
be obtained, but the impurities present in the galena soon

ATER COOUNG

Fig. 12. — Borchers' Cell for refining Lead in Fused Electrolytes.

caused the melting point of the bath to rise above a low red
heat, when the process becomes impracticable.

^Dsrc.

The preparation of metallic zinc has been accomplished
not only by electrolytic processes in aqueous solutions (see
p. 58), and by electrothermal methods (see p. 139), but
also by electrolysis of fused zinc chloride.

The first semi-technical experiments were conducted
by Borchers, who used as electrolytic cell a leaden vessel
with a close-fitting lid hermetically sealed in position by
fused zinc chloride. As anode a vertical carbon rod was
employed, and as cathode a bent piece of strip zinc. Pro-
vision was made for recharging and drawing off the liberated
chlorine gas. Extraneous heat was required to start the
furnace, which was subsequently maintained by the current.
In the Ashcroft-Swinburne process zinc chloride produced by
the action of dry chlorine on blende at 6oo° C. to 700 C,
after treatment with lead to remove the silver and scrap

124 INDUSTRIAL ELECTROMETALLURGY.

zinc to remove the lead followed by solution filtration and
concentration, is fused in enamelled iron pans *> to remove
most of the water, the rest of the water being removed by a
primary-electrolysis between a molten zinc cathode and
carbon anodes, as suggested by Lorenz. 21

The deposition of zinc took place in a firebrick-lined
sheet-iron vessel on the base of which molten zinc acted as
cathode. Carbon anodes were used and a cast-iron gas-
tight roof was employed similar to that used by Borchers.
A slight vacuum was maintained to ensure the removal of
the chlorine.

When sodium chloride was added to the electrolyte in
molecular quantities to the zinc chloride present, a high
current efficiency of 98 per cent, was obtained with a voltage
drop of 4*5 volts per cell at a temperature of 450 C, and
a current of over 3000 amps, or 43 amps, per sq. dcm. of
cathode surface. The decomposition potential of zinc
chloride is 1*49 volts according to Lorenz, 22 whilst Suchy 28
gives 1*57 to i'6o volts. Thus the energy efficiency is
approximately 35 per cent. ; 1000 kw. hours would be
necessary to deposit 260 kgm. zinc. When a high tempe-
rature is used (6oo° C. and over, the m.p. of pure zinc chloride
is 365 C), there is a considerable loss of zinc due both to
the formation of metal cloud in the electrolyte and also to
the volatilization of zinc. This can be much reduced, as
noted above, by the addition of potassium or sodium
chloride, 24 which also serves to increase the conductibility of
the electrolyte. Vogel 25 conducted similar experiments to
those of Ashcroft and Swinburne, using fused zinc chloride as
electrolyte without the addition of any sodium chloride.
He found it impracticable to use a higher current density
than 16 amps, per sq. dcm. with an applied E.M.F. of 45
volts at 450 C.

The disadvantage of these processes is to be found in the
preparation of the fused zinc chloride free from water.
Vogel adopted the expedient of evaporation in vacuo,
whilst, as already indicated, Swinburne removed the last
traces by electrolysis using carbon anodes as an oxygen

ELECTROLYSIS IN FUSED ELECTROLYTES 125

depolarizer. Both methods are exceedingly expensive and
somewhat troublesome.

Snyder 28 suggested that in the direct fusion of blende
with carbon and iron-lime fluxes in a d.c. furnace partial
reduction by electrolytic means takes place, resulting in
the formation of zinc at one electrode and carbon disulphide
at the other. The distinction between electrothermal and
electrolytic reduction is, however, by no means clear in
those cases where carbon is added to the melt.

ALUMINIUM.

The only commercial process for the extraction of alu-
minium from its ores is the thermal electrolytic method
introduced by Hall in America and Heroult on the Conti-
nent in the year 1887. Although aluminium in the form
of complex silicates forms a great portion of the earth's
crust, clays containing some 15 per cent, of aluminium,
yet the economic production of the metal from these
sources is at present an unsolved problem.

The chief raw material is bauxite, obtained in large
quantities from Ireland (I^arne), France (Rhone Valley),
and North America (Alabama), and cryolite obtained from
Greenland.

The composition of bauxite varies with the source ;
the following represent typical analyses : —

Per cent.

Irish.

French.

American.

Austrian

A1 2 3

• • 56

54' 1

Fe 2 3

•• 3

io*4

0.02 • •

. . 12

1*20

Ti0 2

•• 3

—

Water and volatile
matter . . . . 26 12 32 21*9

For the production of pure aluminium the impurities
in the bauxite have first to be removed. There are three
processes pf purification which have received technical
application. In Hall's process (1901) the bauxite is first
calcined mixed with 10 per cent, of carbon, and fused in a

126 INDUSTRIAL ELECTROMETALLURGY

carbon-lined electric furnace. If the iron content is too
low more is added, and the easily reducible impurities
are removed by settling to the bottom as a metallic alloy.
The alumina resulting from the purification of the bauxite
is, however, not so suitable as alumina prepared by the wet
processes, since owing to the high temperature employed
(m.p. A1 2 3 2000 C.) the alumina is prepared in a form
which does not easily dissolve in the electrolyte employed
for the production of aluminium. The addition of metallic
aluminium powder has been suggested for the reduction of
the impurities instead of carbon.

In the Heroult process the crushed bauxite is gently
roasted to remove water and organic matter, then powdered
so as to pass a 30-mesh screen. The powdered material is
digested with caustic soda solution, sp. gr. 1*45, under a
pressure of 6 atmospheres for three hours, during which
period the aluminium passes into solution $s sodium alumi-
nate. After filtration through wood pulp filters into lead-
lined vats, the alumina is reprecipitated by carbon dioxide.
Silica is also thrown down in the process, and since the
alkali is converted into carbonate it has to be recausticized.
Bayer modified this process to overcome these objections
by adding to the sodium aluminate solution some precipi-
tated aluminium hydroxide made in a previous operation,
when, after 36 hours under agitation, about 70 per cent, of
the dissolved aluminium hydroxide can be recovered.
The alumina is washed, dried and finally roasted to about
1100 C. to render it non-hygroscopic, whilst the soda
solution, after concentration in a triple-effect vacuum evapo-
rator, is utilized for extraction of a fresh quantity of bauxite.
Over 40 per cent, of the cost of manufacturing aluminium
is stated to be found in the purification of the bauxite.

The electrolyte consists essentially of a solution of
alumina in fused cryolite (AlF 3 .3NaF), with or without the
addition of a variable amount of sodium fluoride, calcium
fluoride, aluminium fluoride, and occasionally small quan-
tities of the chlorides of sodium or calcium.

In the Hall process the electrolyte is prepared by

ELECTROLYSIS IN FUSED ELECTROLYTES 127

treating a mixture of alumina, cryolite, and fluorspar with
hydrofluoric acid in a lead-lined vat. After drying, the mass
of mixed fluorides is melted in the electrolytic smelting
furnaces.

It is stated that the electrolytes used in the Hall and
Heroult processes have the following components : —

Per cent Halt Per cent Heroult.

A1F 3 . . 590 per cent. AlF^NaF . . 280 per cent.

NaF ... 21-0 CaF 2 *.. ■■ 156

CaF a .. 200 AIF3 564

These electrolytes dissolve some 20 per cent. A1 2 3 at the
temperatures employed.

The Hall furnaces are of cast iron lined with carbon,
and at Lockport, N.Y., are some 1 metre long by 180 cms.

carbon anodes

Fig. 13. — Hall Furnace for the Electrolytic Production of Aluminium.

wide, and 1 metre deep. The carbon liner serves as cathode,
whilst a number of carbon rods 44 sq. cms. in cross-section,
mounted in a special holder, some 40 to the holder and
four holders to each bath, serve as anodes.

The furnaces are worked in series, each anode taking
250 amps. The total current being nearly 10,000 amps,
represents a cathode current density of 100 amps, per sq.
dcm., and at the temperature of working (below 980 C.)
the applied E.M.F. per cell is approximately 5-5 volts.
Aluminium is regularly deposited on the carbon base and

* With an addition of from 3 to 4 per cent, of calcium chloride.

128 INDUSTRIAL ELECTROMETALLURGY

serves as cathode, being in contact with the carbon, whilst
the anodic oxygen liberated by the reaction —

2Al 2 3 ->4Al+30 2

consumes the carbon anodes according to the equation —

A1 2 3 +3C=2A1+3C0

which have to be maintained less than two inches from the
molten aluminium. The furnaces are tapped once a day.
The removal of alumina from the electrolyte is accom-
panied by a rise in voltage across the electrodes, indicated
by the luminescence of a low-voltage lamp shunted across
the bath terminals. Fresh alumina is continuously fed in
to maintain as low a voltage as is convenient.

According to Pring* 7 about one-half of the energy is
expended in the chemical work of decomposing the alumina,
and the remainder is converted into heat which serves to
keep the bath at the proper temperature.

To maintain the temperature the surface of the electro-
lyte, which is usually solid owing to the formation of a thick
crust, is covered with a la} r er of powdered carbon or granu-
lated charcoal. This also serves to obviate the burning
away of the anode electrodes at the point where they enter
the electrolyte by maintaining a reducing atmosphere of
carbon monoxide. Whitewashing the anodes has also been
suggested as a good remedy for this trouble.

The original Hall furnaces were externally heated, but
this method of procedure has now been dispensed with.
Not only is the internal electric heating more economical,
but the iron vessel is protected from attack by the forma-
tion of a crust of electrolyte on the cooler surfaces.

The Heroult furnaces are on similar designs to the Hall,
and are made either round or rectangular in section. Carbon
cathodes in an iron containing vessel are employed ; the
anodes, however, are usually stouter, occasionally up to
35*5 cms. in diameter. Special precautions are taken in
the Heroult design to make use of the protecting crust of
solidified electrolyte.

ELECTROLYSIS IN FUSED ELECTROLYTES 129

Working Temperature. — The working voltage is about
7*0 with a cathode current density of 190 amps, per sq. dcm.
A considerable divergence is found amongst the published
figures for the operating temperature of the cryolite electro-
lytes. The usual temperature is in the neighbourhood of
8oo° C, but temperatures as high as 1000 C. and as low
as 750 C. have been employed.

Cryolite melts at 1000 C. 28 The melting point is first
lowered and then raised by the addition of alumina, as is
indicated by the temperature composition diagram.

(too (

*9o

980

\ /

97o

\ /

OftO

r \ /

^9oo

1 f

Aao

9So

9*o

9oo

q e e to it iq

le \h 20

Fig. 14.— Melting-point composition diagram for alumina

dissolved in cryolite.

To obtain low temperature electrolytes the addition
of other substances is necessary, as has already been men-
tioned. The electrolyte 2(AlF 3 3NaF)3CaF 2 is said to have
an m.p. of 820 C, whilst the addition of the somewhat
volatile sodium chloride lowers the m.p. to under 710 C.
When it is remembered that the m.p. of aluminium is
657 and the b.p. 1800 C., 29 the importance of working at
a low temperature will be obvious.

If too high a cathode density be employed the efficiency
falls off owing to the resolution of aluminium in the electro-
lyte due to the formation of metal fog ; furthermore, the

ii. 9

Specific

gravity.

Solid.

Fused.

2*66

2*54

2*92

2-08

2*90

2'35

296

1-97

2-98

2*14

130 INDUSTRIAL ELECTROMETALLURGY

deposited metal may contain traces of calcium and sodium
formed by electrolysis of the calcium and sodium fluoride
present. The following figures by W. Richards 30 indicate
how closely the molten metal approximates in density to the
electrolyte; when solid the specific gravity of electrolyte
is actually greater than that of the metal : —

Aluminium, commercial . .

V*l V Vyli LC •• •• •• •• •• ••

Cryolite saturated with A1 2 3

Cryolite and aluminium fluoride, AlF 3 3NaF

Cryolite and aluminium fluoride saturated with
alumina

In practice it is found advisable not to add too much
sodium fluoride, since although this lowers the melting point
yet it increases the solubility of the aluminium, and arc
formation may occur. Aluminium fluoride also lowers the
melting point, but it raises the sp. gr. of the melt; its
addition should therefore be controlled. Calcium fluoride S1
appears to be the best addition substance, as it forms a
eutectic at 815 C. with 37 mols. per cent, of A1F 3 .

Current Efficiency. — The actual decomposition volt-
ages of the various salts comprising the electrolyte are not
accurately known. Experiments made by G. Gin and
Minet 32 generally confirm the figures of Richards and Minet
arrived at by calculation.

Decomposition Calculated value

Voltages calculated. Observed assuming complete

Salt. Gin. Richards. value. anodic depolarization.

A1 2 3 . . 279 2*8 2*3 22

C+4F'

AlFg . . 393 40 249 250 j =C f 4+4 @

NaF ... 47 — —

E*(Al)/wAl 2 (S0 4 ) 3 =+i-28 volts.

The current efficiency of a furnace operating at a tempe-
rature of 900 C. is about 65 per cent. An increased efficiency

ELECTROLYSIS IN FUSED ELECTROLYTES 131

results in lowering the temperature owing to the reduction in
the formation of metal fog. At 750 C. a current efficiency
of 95 per cent, has frequently been obtained. It will be
noticed that the energy efficiency of the furnace is low ;
assuming the best working conditions are maintained with
a 95 per cent, current efficiency and 5*5 volts per furnace, the
energy efficiency is only

95 X — =38 per cent.

With 100 per cent, energy efficiency 1000 kw. hours
would produce 153*2 kgm. of metal; in most works the
output is approximately 25 kgm. per 1000 kw. hours.

Anode Consumption. — It has already been noted that
practically complete anodic depolarization is obtained by
the liberated oxygen consuming the anodes, forming carbon
monoxide and with high current densities a mixture of
carbon monoxide and dioxide according to the equations —

Al 2 3 +3C=2Al+3CO
2A1 2 3 +3.C=4A1+3C0 2

When the voltage of the bath is allowed to rise owing to
lack of dissolved alumina, anode effects may occur due to
the liberation of halogens, either fluorine or chlorine if sodium
chloride be present in the electrolyte. Halogen depolariza-
tion is also complete at this temperature, resulting in the
formation of CF 4 or CC1 4 . For every kilogramme of metal
produced the consumption of carbon electrode is roughly
06 kgm. from this cause alone. The electrodes must be
maintained within two inches of the molten metal in order
to reduce the resistance voltage loss over the furnace ;
this can only be accomplished by regulating the distance by
observation of the ammeter and voltmeter. Frequently
internal arcing is caused, accompanied by an increased
electrode loss.

If the anodes are not thickly protected by whitewash
they are occasionally oxidized by the air at the point where
they enter the crust of molten electrolyte, and long pieces
of carbon drop into the bath. These additional losses

132 INDUSTRIAL ELECTROMETALLURGY

bring the electrode loss up to nearly weight for weight
with the aluminium deposited, although with careful work-
ing the former figure of 0*6 kgm. per kgm. metal can be
obtained. The anode carbon must be of high grade to
prevent undue contamination of the aluminium with iron.
Technical working anode current densities vary from 80
amps, per 100 sq. dcm. in the Heroult to over 400 in the
Hall. Blount 33 gives the following analysis of commercial
aluminium, indicating the high degree of purity actually

obtained : —

I. il UL

Al .. .. 99-59 99-00 98-45

Si 025 0*87 1-29

Fe ... ... 016 0*13 o-io

Wright 34 gives the following estimates of costs of pro-
duction per kgm. aluminium : —

1.
Power

Carbon electrodes
Alumina
Labour, repairs, interest

on capital, superin-

tendance

Costs per kgm.

II.

4-8

Power

4-6

4 - 4

Carbon electrodes . .

3*5

8-8

Alumina

I3'9

4'4

Miscellaneous

i"3

22'4

23'3

Other electrolytes have been suggested from time to time,
but have not received technical application ; amongst the
more important may be mentioned AI2S3 in molten cryo-
lite. 35 The advantages gained owing to the low decompo-
sition voltage of the sulphide (0*90 volt) are more than
negatived in practice by the difficulty in preparing the
sulphide from bauxite. Minet 36 used a solution of cryolite
in sodium chloride.

The annual world's output exceeds some 10,000 tons
produced in eleven factories, of which three are in the U.S.A.,
two in France and Great Britain, and one each in Canada,
Switzerland, Austria and Germany.

ELECTROLYSIS IN FUSED ELECTROLYTES 133

Aluminium Alloys.

The earlier experiments by Cowles on the electrothermal
reduction of alumina by means of carbon in the presence
of other metals such as copper led to an extended investiga-
tion of the mechanical and chemical properties of aluminium
alloys. At the present time there is an increasing demand
for a large variety of aluminium containing complexes, and
although the Cowles process, which at one period was
successful on a technical scale, appears to be no longer
in operation, yet it had evident advantages for alloys
containing but small quantities of aluminium. A more
rigid control over the composition and thermal treatment
of the substances is obtained by simple fusion of the re-
quired metals.

Amongst the more important alloys may be mentioned —

Alloy. Percentage composition.

Al. Mg. Cu. Ni. Zn. Sn. Cd.

Gold bronze 3-5 — 97-95 — — — —

Steel bronze 8-5 — 91-5 — — — —

Acid bronze 10 — qo — — — —

Aluminium \ Q g A _. _

bronze J ^ ^

Duralium . . 79 11 10 — — — —

Magnalium 90-98 10-2 — — — — ■ —

Argentum 7 — 70 23 — — —

Rolling alloy 95-5-9 1 — 3-4 *'5-5 — — —

Casting alloy 75 — 62 — 5 12

tST 4 } '■' - 70 - a;- 5 - -

Optical in- \

strument [ 90-5 — — — — 9*5 —

alloy J
"Tiers argent " 66 with 33 per cent, of silver.

The wide application of aluminium alloys for technical
purposes is the subject matter of the VTIIth and IXth
Reports of the Alloys Research Committee of the Institute
of Mechanical Engineering, in which the chemical, physical
and mechanical properties of a very large number of in-
dustrial alloys are dealt with.

134 INDUSTRIAL ELECTROMETALLURGY

REFERENCES TO SECTION II.

1 Electrochem. Ind., 1, 14; 1902.
1 Zeit Elektrochem, 8, 817 ; 1902.
8 Zeit. Elektrochem., 8, 697; 1902.

4 Elektrometailurgie der Alkali Metalle.

5 Trans. Atner. Electrochem. Soc, 9, p. 123, 1906 ; Electrochem. and

Met. Ind., 4, 218; 1906.
• Met. and Chem. Eng. t 8, p. 253 ; 1910.

7 Partington, " The Alkali Industry."

8 Zeit. Elektrochem., 7, 252; 1901.

9 Zeit. Elektrochem., 7, 408; 1901.

10 Trans. Amer. Electrochem. Soc, 17, p. 244 ; 1910.

11 Zeit. Elektrochem., 14, 216; 1908.

12 Zeit. Elektrochem., 11, 612; 1905.

13 Zeit. Elektrochem., 8, 757; 1902.

14 Trans. Amer. Electrochem. Soc., 18, 117; 1910.

15 Zeit. Elektrochem., 81, 612 ; 1905.

18 Journ. Amer. Chem. Soc, 27, 1403; 1905.

17 Zeit. Elektrochem., Nov. 1902.

18 Allmand, "Applied Electrochemistry." 1912.

19 Zeit. Elektrochem., 18, 219; 1907.

80 Electrochem. Ind., 8, 63 ; 1905.

81 Zeit. Anorg. Chem., 89, 389; 1904.
88 Zeit. Anorg. Chem., 12, 272; 1896.
88 Zeit. Anorg. Chem., 27, 152 ; 1905.

84 Griinauer, Zeit. Anorg. Chem., 89, 389; 1904.
25 Trans. Farad. Soc, 2, 56; 1906.
88 Electrochem. Ind., 4, p. 152, 1905.

27 " Some Electrochemical Centres." 1908.

28 Pyne, Trans. Amer. Electrochem. Soc, 10, 63 ; 1906.

29 Greenwood, "Electrochem. and Metal. Ind.," p. 408, 1909.

30 Zeit. Elektrochem., 1, 307; 1895.
81 J. S.C.I. , 367 ; 1913.

32 V. Int. Congress Applied Chemistry.

33 " Practical Electrochemistry." 1901.
84 "Electric Furnaces." 1904.

36 G. Gin, D.R.P. 148627 of 1908.

86 Borchers, " Elektrometailurgie," p. 108. 1905.

BIBLIOGRAPHY.

" ElektrolyseGeschmolzenerSalze." R. Lorenz. Knapp. Halle. 1905
" Principles of Applied Electrochemistry." Allmand.
4 * The Alkali Industry." J. R. Partington. Bailliere, Tindall & Cox.
" Die Gewinnung des Aluminiums." A. Minet. Knapp. Halle. 1902.
" Elektrometailurgie der Alkalimetalle." H. Becker. 1903.

Section III.— THE ELECTROLYTIC
PREPARATION OF THE RARER METALS

Gaiaxum.

Gaujum is conveniently deposited on a platinum cathode
from the complex gallate formed on solution of a gallium
salt in excess caustic soda. The deposit can be melted off
the cathode under warm water (m.p. 30*15°, but can be
supercooled to o° C).

Indium.

According to Schucht l the neutral sulphate is the most
suitable electrolyte to use. Dennis and Geer 2 suggest
the nitrate or chloride, with the addition of a reducing
agent such as formic acid. Thiel 3 suggests a weakly acid
bath containing sulphuric acid and ammonium sulphate.
The electrolytic potential of indium is approximately
Ea=+o*45 volt, and resembles cadmium.

Thaujxtm,

I,epi£me in 1893 suggested the double oxalate of
ammonium and thallium as a suitable electrolyte. Forster
made use of a neutral sulphate electrolyte, deposition
taking place on a copper cathode of 100 sq. dcm., using
a platinum anode of 8 sq. cm. and a current of 1*3 to 1*5
amps, at 35 volts. The metal can be fused under KCN.
Partial precipitation on the anode as T1 2 3 is liable to occur,
especially in the presence of reducing agents such as acetone.

The electrolytic potential of thallium is approximately
E*= +0322 volt (for Tl/Tl' solutions), the metal thus
resembling cobalt or iron. The reducing power of the

136 INDUSTRIAL ELECTROMETALLURGY
thallium salts in terms of the electrolytic potential difference

Tryrr is—

tt*

E A =I-I99+0024 log, -j^r volts.

Fused Electrolytes. — The elements cerium, neody-
mium, praseodymium, lanthanum and samarium are
most conveniently prepared by electrolysis of the fused
anhydrous chlorides. They are all white metals with a
slightly yellowish tinge and fairly stable in air, lanthanum
being the most easily oxidized.

The temperature necessary for electrolysis varies for
each metal, as seen from the following table : —

Ce .
La
Nd ,
Pr .
Sm

Hildebrand and Norton advised the use of iron electrodes ;
the cathode being placed in a porous porcelain cell contain-
ing the fused chloride, protected by a layer of ammonium
chloride. In the anode compartment surrounding the
porcelain cell a mixture of fused sodium and potassium
chloride was used.

Muthmann advocated the use of a water-cooled copper
electrolytic cell containing two vertically situated carbon
electrodes. As electrolyte he used the fused chlorides,
with or without the addition of the chlorides of sodium,
potassium and barium. He recommends the following
electrolyte for cerium : —

Fusion point

M.p. metal.

of chlorides.

625° c.

—

8io°

907°

840°

785

940°

818°

i300°-i400°

686°

CeCl 2

Na.Cl.KCl

BaCl 2

200 parts.
15-20 parts.
A trace.

Electrolysis with a current of 120 amps, at 12-15 volts
yielded 750 gms. of metallic cerium in 6 hours.

With samarium a very high cathode current density is

PREPARATION OF THE RARER METALS 137

required to ensure the fusion of the metal. The addition
of J part by weight of barium chloride to the chloride is
advised.

Boron.

Experiments by Hampe 4 on the electrolysis of molten
borax indicated the formation of a sodium boron alloy
at the cathode. Lyons and Broadrill, 6 using a fused borate
electrolyte and a carbon anode, claim the preparation of
boron by reduction of the B 2 3 .

The carbides B 2 C 2 , B 6 C, are the products of electro-
thermal reduction (see p. 172).

Vanadium.

M. Gin 6 suggested the electrolysis of molten vanadium
fluoride between an iron cathode and a compressed mixture
of carbon and vanadium trioxide as anode material in a
cell lined with alumina. With a cathode current density
of 600 amps, per dcm. and an anodic one of 200 amps,
per dcm. and an E.M.F. of 11-12 volts, pure vanadium could
be deposited on the cathode with the reformation of vanadium
fluoride at the anode —

2VF 3 =2V+3F 2
3^2+ V 2 3 +3C=2VF 3 +3CO

An alternative method is the use of a carbon anode in an
electrolyte of V 2 3 dissolved in a double fluoride, 2VF 3 .3CaF 2 .
Wood's process entails the use of the oxide and calcium
oxide as electrolyte, requiring a much higher temperature.

Titanium.

Borchers 7 patented the use of calcium chloride as
electrolyte with the continuous addition of titanium dioxide
for the preparation of the pure metal.

Pederson 8 suggests copper titanium as an industrial
alloy suitable for many purposes; it is prepared by the
electrolysis of titanium dioxide in calcium fluoride as
electrolyte, using a copper cathode.

138 INDUSTRIAL ELECTROMETALLURGY

Manganese.

Experiments by Guntz 9 and Bunsen 10 on the electrolysis
of concentrated solutions of manganous chloride indicated
that the preparation of the metal free from all traces of
oxide was a matter of great difficulty. Better results are
obtained by electrolysis of the fused chloride or fluoride
in an alkali chloride electrolyte. Simon 11 suggests calcium
fluoride as electrolyte, adding manganese oxide continuously
to the electrolyte in a manner similar to that adopted in
the production of aluminium.

Uranium.

This can most conveniently be prepared by electrothermal
methods (see p. 153), but very pure metal can be deposited
by electrolysis of the fused chloride UC1 4 between carbon
electrodes. 12

REFERENCES TO SECTION IIL

1 Berg. U. Hutten. Zt., 39; 1880.

2 Ber., 37, 961 ; 1904.

3 Zeit. Anorg. Chem. 40, 280; 1904.

4 Chem. Zeit. , 12, 841.

6 U.S. Pat. 785962, 1905.

6 Int. Cong., Appl. Chem., 1903.

' D.R.P. 150557.

8 Elektrochem. Zeit., April, 1914.

9 Bull. Soc. Chem., 3, 275 ; 1892.

10 Pogg. Ann., 91, 619; 1854.

11 Eng. Pat. 17190, 1900.

18 Rideal, " Das Elektrochemische Verhalten des Urans." Diss.
Bonn, 191 3.

BIBLIOGRAPHY.

" Die Darstellung des Chromes und Seiner Verbindungen." W. Le
Blanc. Halle. 1902.

Section IV.— ELECTROTHERMAL PROCESSES

Zinc.

We have already referred to the electrolytic deposition of zinc
in both aqueous and in fused solutions, but the most serious
rival to the ordinary Belgian thermal practice is to be found
in the electrothermal processes. The ordinary method of
smelting zinc suffers from serious disadvantages. In general
practice the ore, after roasting to convert the sulphide or
carbonate into the oxide, is mixed with about half its weight
of coal slack or coke, and heated in small fireclay retorts.
Owing to the high temperature necessary to expel the zinc
(over iioo° C.) the retorts must be small, holding only some
30-40 kgms. of the charge : the distillation of the zinc is
completed in 20 hours. With a high-grade ore one and a half
tons of coal per ton of ore is the minimum consumption,
whilst even 4 tons of coal per ton of ore may be required
in a badly operated furnace with a low-grade ore. The life
of a retort is short, averaging from 30 to 40 days, being
attacked not only by the hot gases outside and the zinc
vapour inside, but also by the slags, especially by those with
a high lime or iron content. J . W. Richards l has calculated
the thermal efficiency of the average furnace to be under
7 per cent.* The process is further complicated by the
difficulty of removal of the infusible slags remaining in the
retorts.

Not only are the furnace operation costs high in both
labour and material, but great difficulties are met with in

* For information on the thermal conductivities of various furnace
liners, as well as the heat loss from furnaces and electrodes of different
shapes and sizes, see Northrup, McLeod, Kanolt, Fitzgerald, Langmuir, and
others in Trans. Atner. Electrochem. Soc, 1912 to 191 7 ; also Bronn,
" Der Elektrische Of en."

140 INDUSTRIAL ELECTROMETALLURGY

the condensation of the zinc vapour. There is a substantial
loss due to diffusion of the vapour through the walls of the
retort and to the retention of part of the zinc in the slag,
especially if the sulphur has not been entirely eliminated
by roasting. Again, in the actual process of condensation
of the zinc only a part coalesces to a regulus " spelter," the
remainder being recovered as " blue powder."

The formation of " blue powder " is more common in
electrothermal processes than in the Belgian, but is by no
means an unimportant factor in the latter.

There are three factors which are considered to have
an influence on the formation of " blue powder " :

I. The formation of an electrostatic charge on the zinc
vapour globules daring the process of condensation.

II. The rapid chilling of the zinc globules in the condenser
(the m.p. of the metal being 419 C). Rapid cooling to,
say, 400 C. may considerably undercool the globules before
they are run together. Dilute zinc vapour is more liable
to be undercooled than more concentrated ones. The
optimum condensing temperature has been found to lie
between 500 C. and 850 C, depending entirely upon the
concentration of the issuing vapour.

III. Superficial coating of the condensing globules with
an oxide skin. This factor is probably the most important
where very low spelter recoveries are made. Reduction
of the zinc oxide may take place according to either of the
following equations : —

(i) ZnO+C$Zn+CO
(ii) ZnO+CO$Zn+C0 2

The main reaction following that indicated in the second
equation. Owing to the fact that reduction does not proceed
with sufficient rapidity under noo° C, the reduced metal is
not removed from the sphere of action by condensation, as
is the case with most metals, e.g. iron or copper, but remains
in the gaseous phase. The reaction is consequently revers-
ible and partial reoxidation of the reduced zinc may take
place, especially if the free space between the packed charge

ELECTROTHERMAL PROCESSES 141

and the condenser be too great. The formation of carbon
dioxide is usually reduced to as small an extent as possible
by addition of excess carbon to the charge, when the result-
ing gas expelled with the zinc vapour will consist chiefly
of carbon monoxide containing but little of the dioxide.
The exact ratio CO : C0 2 will depend on the temperature
of operation, being governed by the equilibrium —

(iii) 2CO$C+C0 2

The following figures indicate the composition of the equi-
librium gas mixtures at various temperatures : —

Temp. °C.

Per ceat. C0 2 .

Per cent.. CO.

450

550

89-3

107

650

750

850

950

i*5

98-5

1050

034

99-6

Other gases, such as oxygen, water vapour, hydrocarbons or
silicious dust, may all assist in the formation of a film on the
condensing zinc globules.

J. Johnson 2 gives the following figures for the vapour
pressure of zinc at different temperatures : —

Vapour pressure
Temp. in mm. of mercury.

920 C. 750

750 100

700 50

610 10

500 I

420 IO" 1

419 m.p.
350 10-2

290 io -3

It will be noted that the vapour pressure of the zinc only
becomes small when the gas is cooled to 600 ° C. Even at this
temperature the gas can contain 1*3 per cent, volume of
zinc vapour without deposition of any metal, while at the

142 INDUSTRIAL ELECTROMETALLURGY

same time over 75 per cent, of the original carbon monoxide
has been converted to the dioxide. The effective condensation
of zinc can, therefore, never be complete ; " bine powder "
is always formed, but the quantity can be reduced by the
production of a gas rich in zinc vapour and providing a very
rapid fall in temperature from 1100 C. to between 6oo° C.
and 700 C. in a very short space. Under these conditions
advantage is taken of the relative slowness with which equi-
librium will be re-established by cooling to this relatively
low temperature according to equations (ii) and (iii).

Instead of redistillation of the " blue powder " alternative
treatment by electrolysis in fused or aqueous solution as
suggested on p. 59 might prove practicable. If anodic
depolarization by means of the free zinc in " the blue
powder " (briquetted to anodes) could be made use of, the
cost of electrolyte recovery would be reduced to the operation
of a refining process.

Power Consumption. — Harbord 3 gives the following
figures obtained in test runs at Trollhatten, working with
a blende calamine mixture (30 parts Broken Hill ore, 1 part
calamine, and 7*5 parts coke dust) ; the blue powder (con-
taining 54 per cent, zinc and 20 per cent, lead) and oxide
recovered from this charge was mixed with a further quantity
of blende coke dust and lime, and distilled in a second
furnace.

Energy consumption per Electrode

1000 kgm. of ore smelted. consumption.

2078 kw. hours. 31*5 kgm.

The above figures include the necessary energy for redis-
tillation of the blue powder. Mounden 4 estimates the
recovery in these works to be 75 per cent, of the zinc, 80 per
cent, of the lead, and 80 per cent, of the silver. Salgues at
Artege, in France, using a 40-45 per cent, zinc ore, obtained
1000 kgm. zinc with a current consumption of 4800 kw.
hours, or per 1000 kgm. of ore smelted 2016 kw. hours
were required. G. Gin 5 calculates the current of energy
required for smelting 1000 kgm. of ore containing 50 per
cent, zinc at 1500 kw. hours, while according to Stansfield 6

ELECTROTHERMAL PROCESSES 143

Snyder has smelted pure zinc oxide with an energy con-
sumption of 1050 kw. hours per 1000 kgm. of oxide.

Harbord's figure includes the electrical energy con-
sumption necessary for the redistillation of the blue powder,
being about 500 to 600 kw. hours per ton of blue powder.

We may take the average power consumption per ton
of ore at 1500 kw. hours, as opposed to the maximum and
minimum coal consumption of 4 and ij tons per ton of ore
used in the Belgian process.

Under normal working conditions the electrode loss is
estimated at 4^. to 6d. per ton of ore used, and is thus less
costly than the retort consumption of 8d. per ton in the
Belgian process.

Johnson 7 estimates the electrode consumption at from
1 to 15 kgm. per ton of ore, figures considerably under those
obtained by Harbord. It is evident that cheap power
rates are essential to the successful operation of electro-
thermal zinc smelting process.

Types of Zinc Furnaces employed. — The chief advan-
tages to be gained by electrothermal smelting processes
is the feasibility of working with charges larger than with the
Belgian retorts, and the possibility of continuous operation.
Furnaces taking two to three tons per charge have proved
satisfactory, while Johnson 8 does not contemplate serious
difficulties in operating 10-ton capacity units. Other
advantages which become increasingly important when a
low-grade zinc ore is used are the possibilities of providing
an easily fusible slag which may be tapped off and worked
up for valuable metals such as silver and copper, while
under certain conditions a molten metal may be run o£E
(especially in ores with a relatively high lead content) in
addition to and separate from the slag : conditions scarcely
possible in small retorts.

The slag fusion temperature should be adjusted to lie
just above the temperature necessary for distillation of the
zinc to avoid inclusion of the metal. In silicious slags the
addition of silica is the controlling factor, in basic slags
carbon. 9

144 INDUSTRIAL ELECTROMETALLURGY

Resistance Furnaces. — The original application of
electrical heating to zinc smelting was made by Cowles in
1880, who adopted a simple form of resistance furnace. ' In
more modern form resistance furnaces designed by Johnson 10
are in use in the United States, and by Salgues n at use in
Pyrenees and Trollhatten (Sweden, 20,000 kw.). Sarpsborg
{3000 kw.) and Hallstahammer in Norway.

The Cowles and the earlier Johnson furnaces were
operated with horizontal electrodes inserted in the ends of
an arched chamber of firebrick lined with a refractory,
such as fireclay or bauxite. The charge containing ore

Fig. 15. — Resistance Zinc Furnace. Johnson type.

and coke was used as resistance, and the furnaces were
intermittent in action. The later forms of the Johnson
furnace, as well as those of Salgues, have vertical electrodes
and are continuous in operation.

The roasted ore, mixed with carbon and lime or other
flux, is fed in through the hopper A into the smelting chamber
J, in which are situated three electrodes B, C, and D. The
lower electrode D, a carbon plate, is usually covered with
molten lead containing silver (E) when ores containing lead
are used ; above this is a layer of molten slag F. These two
layers have separate tapping holes. The zinc vapour
together with a mixture of carbon dioxide and monoxide
is passed off to the condenser through the column I, which

ELECTROTHERMAL PROCESSES

145

is filled with broken carbon maintained at 1100 C. By
this means pure zinc vapour with carbon monoxide as the
only diluent is produced and rapidly condensed in the air-
cooled receiver. Johnson has claimed an 80 per cent, zinc
and a 60 per cent, spelter recovery from a 30 per cent, ore
with this type of furnace.

By introducing the fresh charge under the surface of the
slag the production of smoke is said to be minimized, result-
ing in a decrease of blue powder formation. 12

Johnson gives 13 the following compositions of the slag

matte tapped :-

Slag analysis.

Matte analysis.

Si0 2 . .

. 40

Fe .. ..45

CaO ..

. 22

Cu . . . . 25

MgO . .

S . . . . 29

FeO ..

. 10

MnO . .

AI2O3 . .

ZnO

015

0-05

0'3 oz

./ton.

For low-grade zinc ores containing relatively large quantities
of lead, copper, gold and silver the process offers distinct
advantages. The preheating of the charge before intro-
duction into the electric furnace by the gas liberated effects
a considerable economy. u The resistance furnaces employed
at Trolhatten hold each about three tons of charge, and can
smelt 2 # 8 metric tons of ore per 24 hours. A current of
2600 amps, at about 100 volts is used to operate each
furnace, corresponding to a current density of 128 amps,
per sq. dcm. Two tons of blue powder are re-smelted with
every ton of fresh ore. Various modifications have been
suggested for the treatment of sulphide ores to overcome
the difficulty of the complete removal of the sulphur by
prolonged roasting before the reduction. This difficulty
can be avoided by the additions of a suitable flux which will
remove the sulphur in the slag, such as iron or lime. Snyder 16
i/. 10

146 INDUSTRIAL ELECTROMETALLURGY

suggests treatment of the ore unroasted with iron and with
lime fluxes and carbon in a resistance furnace, with the
simultaneous production of zinc and carbon disulphide.
Brown and Oesterle 16 further improved upon this patent
by claiming the simultaneous production of zinc, carbon
disulphide and calcium carbide.

The C6te Pierron process 17 uses scrap iron to produce
ferrous sulphide according to the reversible equation —

Fe+ZnS^FeS+Zn

the equilibrium being shifted over entirely to the right
through the volatilization of the zinc out of the liquid phase.

The process is suitable for lead-zinc ores, since the lead
can be directly recovered, and the zinc vapour is not diluted
with any carbon monoxide. Against these advantages must
be set the cost of the scrap iron necessary for reduction,
900 kgm. of iron being required for every 1000 kgm. of
zinc and 300 kgm. for every 1000 kg. of lead.

The following costs of production are entailed at Ugine,
Savoy. 18 Cost per ton of zinc produced : Power 10s. 6d.,
depreciation 9s. 6d., electrodes 4$. 6d., iron 3s. 4^., labour
6s. 6d., miscellaneous 4s. 2d. ; total 38s. 6d. Eleven per
cent, of zinc is lost in the process.

In the Imbert-Fitzgerald furnace 19 wedge-shaped carbon
rods are used as a permanent resister for the furnace. A
mixture of one part of ferric oxide and three parts of iron
sulphide are used as a flux at 1100 C, to which six parts
of blende are added ; on the addition of molten copper or
pig iron the zinc is volatilized off and condensed. The copper,
of course, would be recovered from the resulting sulphide
in the usual manner, but in practice iron is used. The
furnace must naturally be worked in a reducing atmosphere.
Dorsemagen suggested the use of a resistance furnace for
the production of zinc and carborundum by the reduction
of siliceous zinc ores, while Borchers patented a process for
the simultaneous production of ferro-silicon and zinc.

The majority of these modified processes in which by-
products from the sulphur as carbon disulphide or silica as

ELECTROTHERMAL PROCESSES

147

carborundum or ferrostticon are obtained have not been
worked on a sufficiently large scale to enable an assessment
of their technical utility to be made.

Radiation Furnaces. — The most important radiation
furnace employed for zinc smelting is that of C. de Laval, 80
which has been used in the United States and also at
Trolhatten.

The charge of roasted ore, coke and flux enters through
the shaft D, where it is exposed to the radiation from the
arc between the horizontal
electrodes A. The CO and D

zinc vapours leave at a high
temperature through B to
the condenser, while the slag
can be removed through
the tapping hole C. It will
be noted that the arc is
operated in a reducing
atmosphere which consider-
ably lessens the electrode
consumption, although 40
kgm. of electrodes per metric ton of ore are required, a
figure considerably higher than given for resistance furnace
operations.

Reduction proceeds quietly, and very pure zinc can be
obtained in these furnaces, while very little of the metal is
retained in the slag. Owing to the low diathermacy of the
ore the electrical efficiency is poor, consuming about 70 per
cent, more power than the resistance type of furnace. The
furnace is very simple to operate.

Copper.

Very few large-scale experiments have been made in the
electrical smelting of copper ore, but some have shown
promising results. The treatment for complex copper ores
may be roughly divided into three classes :

(A) Ores containing metallic copper (e.g. native copper)

148 INDUSTRIAL ELECTROMETALLURGY

can be smelted, directly separating the metal from the
gangne.

(B) Ores containing copper sulphide or arsenide in
addition to iron can be smelted in a blast furnace in which
part of the sulphur in the ore is oxidized, the heat of oxida-
tion assisting in the fusion of the ore. The resulting slag
should contain the oxides and silica as well as most of the
iron, while the matte contains the bulk of the copper as
sulphide or arsenide with a small quantity of iron.

(C) Oxidized ores can be selectively reduced with carbon.
By careful adjustment of the carbon content in the charge
most of the iron can be retained in the slag in an oxidized
condition.

Experiments made at La Praz and Iyivet in France from
1903 to 1907 on the production of copper matte from a
sulphide ore in resistance furnaces, were favourably reported
on by M. Vattier for the Chilian Government.

The furnace used was a simple resistance one of the
Keller type furnished with two pairs of electrodes in separate
zones of the furnace chamber, so that the ore fused in one
zone could be maintained at the fusion point in the second to
effect the separation of slag from matte.

Twenty-four tons per twenty-four hours of ore could be
treated in a furnace of 2 cubic metres capacity.

The power consumption was 500 kw. hours per 1000
kgm. ore and 5 kgm. electrode material.

Vattier gives the following percentage analysis of ore,
charge, slag and matte : —

Charge.

Matte.

Slag.

Cu ..

- 510

47*9

0*10

Fe ..

... 2850

243

32-50

Mil...

. . 7-64

i"4

823

S ..

4 - I2

22*96

0'57

Al 2 O s

4*00

o-5

520

CaO

-. 7*30

9-90

Si0 2

. . 2370

2720

The power consumption for the furnaces averaged 4750 amps.
at 119 volts with a power factor of 0*9.

ELECTROTHERMAL PROCESSES 149

He calculated that 3*2 metric tons of coke (costing in
Chili 3613) were required for the ordinary coke furnaces to
produce 1 metric ton of copper. The same results could be
obtained with 8000 kw. hours of electrical energy. Taking
a figure as high as o'id. per kw. hour, produced by water-
power, the power costs would only amount to £3 6s., while
the electrode consumption and furnace depreciation would
not amount to more than £1 16s. per ton of metal produced.
Under the conditions where fuel costs are remarkably high,
and where electric energy could be produced at very reason-
able rates, electrical processes are clearly indicated.

The electrothermal method of copper smelting has been
tested both at Kaafjord and Trondjhem, Norway, with
success. 21 Experiments in Germany 22 on the reduction
of a silicate ore with calcium carbonate and coke at 1600 C,
obtained a minimum power consumption of 1100 kw. hours
per metric ton, necessitating very cheap water-power. The
preparation of copper nickel alloys in an electric furnace has
been experimented with in Norway. The process should
proceed smoothly owing to the complete miscibility of the
metals in each other, forming solid solutions, as indicated
in the curve on p. 41 . 2S Copper thus prepared is likely to
contain cuprous oxide unless a reducing atmosphere is
continually maintained in the furnace. Heyn, 24 who had
investigated the solubility of Cu 2 in metallic copper, finds
a eutectic containing 3-5 per cent. Cu 2 melting some 25 C.
below the m.p. of the pure metal.

Nicked.

W. Iy. Morrison 26 and S. B. Ladd 26 have described
the conditions necessary for the satisfactory smelting of
oxidized nickel ores. A small furnace has been worked
at Sault St. Marie Ont., U.S.A., while the Consolidated
Nickel Co. at Webster have successfully operated on a large
scale the reduction of a hydrated magnesium nickel silicate
complex containing less than 2 per cent, of nickel.

A resistance furnace of simple type is employed with a

»»
ft
ft

150 INDUSTRIAL ELECTROMETALLURGY

carbon hearth and one or a series of vertical electrodes
entering through the roof. The ore after crushing is mixed
with broken coke, yielding on reduction a nickel ferro-
silicon metal of the following composition : —

Ni . . . . . . 14 per cent.

JC c •• . .. •• •• o

Ol •• •• •• •• aO

Other metals . . . . 2

and a slag consisting chiefly of aluminium and magnesium
silicate containing about 0*5 per cent, of nickel. The power
consumption is about 1200 kw. hours per 1000 kgm. of
ore smelted. F. Clergue has suggested the use of a revolving
electric radiation furnace for the production of ferro-nickel ;
his process is said to be in operation at Essen, Germany (see
also p. 228).

Manganese.

Manganese is usually produced in the forms of spiegeleisen
and ferro-manganese, the demand for the pure metal being
limited. Although it can be prepared in a pure form by
electrolytic methods (p. 138), the electrothermal processes
are quicker and more convenient.

Moissan 27 effected the reduction of Mn0 2 by carbon in
a small arc furnace with a current of 150 amps, at 60 volts,
preparing several hundred gms. of the metal in a few
minutes. He attempted to remove the excess carbon
present in the metal by refusion with Mn0 2 . Borchers M
cpuld not confirm the removal of the excess carbon by this
method. Gin 29 used a mixture of Mn0 2 with sodium
sulphate and carbon in a small arc furnace. By this means
sodium manganate is produced which has a melting point
under 2000 C, from which the manganese can be produced
at a temperature well below its point of vaporization (m.p.

1247 c.).

Tin.

Harden 3° has given details of the conditions necessary
for the reduction of tin ores. Although electrothermal

ELECTROTHERMAL PROCESSES 151

smelting of tin has not been accomplished on a technical
scale, with the exception of tin dross smelting in tin plate
works, 31 yet, owing to the unsatisfactory working of the
ordinary blast furnace where losses by volatilization of
stannic oxide are by no means inconsiderable, the electro-
thermal methods have some prospect of future development.

Chromium.

Metallic chromium is only prepared on a comparatively

small industrial scale, the chief electric furnace production

being ferro-chromium (see p. 234). It can be obtained with

a simple arc furnace using intermittent charging, the fused

metal produced by reduction being broken out. Reduction

is usually accomplished by means of carbon according to

the equation —

Cr 2 8 +3C=2Cr+3CO

The reaction commences at 1185 C. 32 The resulting grey
metal usually contains the extremely hard carbide, Cr 3 C 2 ,
which is difficult to remove.

Refusion with the calculated amount of chromic oxide
usually entails the presence of both oxygen and carbon in
the metal. More effectual removal can be accomplished
by the addition of lime to the charge —

3Cr 3 C 2 +2CaO =9Cr +2CaC 2 +2CO

although small quantities of calcium chromite are formed
under these conditions. Aschermann at Cassel successfully
developed a process for the preparation of chromium by
reduction with antimony sulphide in a small graphite
crucible —

2Cr 2 3 +Sb2S 3 =4Cr +2Sb +3S0 2

The antimony is entirely removed by reheating. Becket 33
uses silicon as a reducing agent —

2Cr 2 O s +3Si =4Cr +3Si0 2

152 INDUSTRIAL ELECTROMETALLURGY

Molybdenum.

Metallic molybdenum, for which there is an increasing
demand in the production of special steels, is more easily
prepared than chromium by the reduction of the oxide
with carbon. A small deficit of carbon according to the
equation —

Mo0 2 +2C =Mo +2CO

ensures the presence of excess oxide in the metal. The oxide
is sufficiently volatile to be easily removed by the sublimation
from the melt. The most common form of molybdenum ore
is the sulphide, and the direct preparation of the metal from
molybdenite is the subject of many patents. Guichard 34
and !Lehner *■ suggested the reduction with carbon in the
presence of lim<

MoS 2 +2CaO+2C=Mo+2CaS+2CO

Becket 36 has claimed the process for reduction with a smaller
amount of carbon than indicated by the above equation —

2MoS 2 +2CaO +3C =2Mo +2CaS +CS 2 +2CO

Calcium carbonate may, of course, be used instead of lime —
2MoS 2 +2CaC0 8 +5C=2Mo+2CaS+CS 2 -f6CO

The addition of calcium fluoride as a flux causes the reaction
to proceed more smoothly. 37 Small traces of iron present
in the molybdenite are removed by volatilization on further
fusion of the metal. Neumann 38 has suggested the reduction
by means of silicon ; according to Keeney 39 —

MoS 2 +Si=Mo+SiS 2

unsatisfactory results were obtained. Calcium carbide has
a growing market as a reducing agent, and is especially
effective for the preparation of metals like molybdenum —

5MoS 2 +2CaC 2 =5Mo +2CaS +4CS 2

ELECTROTHERMAL PROCESSES 153

Tungsten.

This metal is also being used in increasing quantities for
the preparation of special steels and in the electric lighting
industries. For most steel work the metal is usually not
isolated, but reduced to produce ferro-tungsten (see p. 232).
Owing, however, to the variable carbon contents of the
ferro alloy pure tungsten is used for high-grade steel. In the
manufacture of tungsten for steel work and electric lamp
filaments the oxide is usually reduced by means of hydrogen
in an electric-resistance furnace and subsequently melted
to prepare the ductile metal. 40 Metal containing a variable
amount of carbon as carbide and free carbon can be prepared
in the arc furnace by methods similar to those used for the
preparation of chromium.

Vanadium, Titanium and Uranium.

These three elements can be prepared by reduction
with carbon of their respective oxides, V 2 Os, Ti0 2 and U 3 8 .
The resulting metals always contain small quantities of the
carbides and nitrides.

The industrial demand is in the form of the ferro-alloys,
and they are always produced as such.

Zirconium.

According to Moissan 41 this element can be produced
by reduction with carbon with a current of 1000 amperes
at 40 volts in a simple arc furnace. Greenwood 42 found
that no reduction took place below 1400 C.

The element is not produced industrially.

SnjcoN.

There is a limited but growing demand for this element
as distinct from the ferrosilicon alloy for reduction purposes.
Its heat of oxidation, being 215,692 calories per gram mole-
cule, is only exceeded by " thermite " and the alkali metals.

Crude silicon is prepared from silica by reduction with

154 INDUSTRIAL ELECTROMETALLURGY

carbon, and in this state it contains Si0 2 , N 2 , and other
impurities. An effective method of purification 43 is to
treat the crude material for two hours in a crucible covered
with coke, then stir in §-3 per cent, of magnesium powder.
A slag of magnesium silicate is separated, and the silicon
can be poured off into sand moulds. Reduction com-
mences at 1460 C, the melting point of the metal being
1430 C. The Acheson Carborundum works 44 use carbon-
lined firebrick furnaces with two depending electrodes,
the current passing from the electrodes to the hearth.
The furnace is operated as a resistance furnace, since the
element is volatilized at the temperature of the arc (b.p.
2800 C). Each furnace uses 1000 kw., and from 250 to
350 kgm. of silicon can be tapped off every few hours.

According to Stansfield 46 a high-grade unrefined silicon
had the following composition : —

Si 9571 per cent.

Fe
Al
P
C

2 24

1 96
001
008

Potter 48 has suggested the use of silicon carbide as
a reducing agent —

Si0 2 +2SiC=3Si+2CO

Attempts have been made to prepare silicon electrolyti-
cally, notably by Deyille, Minet and Grosz, using as electro-
lyte either sodium potassium silicate or potassium silicate,
adding silica from time to time ; indifferent results were ob-
tained.

A great variety of compounds have been prepared by
the interaction of silica and carbon in the electric furnace,
some of which have become extremely important in technical
work. These will be referred to in a later section (p. 164).

Graphite.

Carbon can exist in at least three well-known modifica-
tions, two crystalline 47 and one amorphous : diamond,

ELECTROTHERMAL PROCESSES 155

graphite and ordinary carbon. The diamond is the stable
modification at low temperatures, whilst graphite is the
stable form above 500 C.

The technical transformation of anthracite, coal or
coke into graphite was first developed by E. Acheson. He
found that the direct conversion of pure carbon into graphite
was a very tedious operation, but that the presence of small
quantities of impurities, especially metals such as iron or
aluminium and certain non-metals such as silicon and
boron, catalytically hastened the conversion.

According to Townsend preliminary ionization is neces-
sary for the formation of graphite from carbon, and the
function of the catalytic material apparently serves to
produce graphite by the decomposition of a carbide formed
by the catalyst with the carbon.

It is assumed that the formation of the carbide takes
place in the hottest zone of the furnace, and as the tempe-
rature is gradually raised the carbide is decomposed leaving
behind graphite, whilst the catalyst is volatilized to the
colder zones, there to recommence the conversion of carbon
to graphite.

That the presence of a catalyst is not absolutely necessary
is shown by the experiments of Acheson, Borchers and others,
but for technical production it cannot be dispensed with.

It has already been noted that graphite is the stable
form of carbon above 500 C, consequently the vapour
pressures of carbon vapour above the solid carbon and
graphite at, say, 1100 C. will not be the same, the carbon
possessing the higher vapour pressure. In the presence
of graphite at 1100 C. carbon will thus gradually sublime
and be redeposited in the form of graphite.

Fitzgerald and Forssell have attempted to measure
the relative vapour pressures of carbon and graphite at
low temperatures between 500 and 700 C. by studying
the equilibrium composition C+C0 2 ^t2CO, when carbon
or graphite in the solid state is present. They found that
at 500 C. the vapour pressure of carbon was 37 times that
of graphite, and at 640 C. 5*4 times as great. .

156 INDUSTRIAL ELECTROMETALLURGY

At Niagara, anthracite is used as the carbon for con-
version into graphite. The furnace consists of a long trough
holding about 6 tons of anthracite mixed with 3 per cent,
of oxide of iron, and is finely crushed to the size of rice
grains. The anthracite surrounds a carbon electrode core
which carries the heating current. Each furnace is about
30 feet long and 2 feet 6 inches wide and deep, constructed
of fireclay bricks with a carborundum slab liner. The
terminal plates at each end of the furnace are water-cooled,
since they have to carry over 15,000 amperes.

a a

carbon for graph ih*ing
St.. thermo couples

Fig. 17. — Resistance Furnace for the production of Graphite.

About 1600 kw. are consumed per furnace, commenc-
ing at 8000 amperes at 200 volts, and as the resistance
decreases with elevation of the temperature the current at
the end of the operation is about 20,000 amperes at 0'8o volt.
The furnace takes about a day to heat up, and from four to five
days to cool down. The resulting graphite is remarkably
pure, usually containing only from o*i to o*8 per cent, mineral
ash, chiefly iron which has not been completely removed by
volatilization.

The adequate protection of the pyrometer couples
embedded in the anthracite is a matter of considerable
difficulty. The limits of the graphitizing zone, which is
well over 2000 C, have to be continuously observed so as to
ensure the presence of a non-graphitic colder anthracite liner
between the graphite and the fusible firebrick.

ELECTROTHERMAL PROCESSES 157

Small traces of sulphurous gases are liberated as well as
carbon monoxide and dioxide during the primary heating
up, which in time destroy metal pyrometer sheaths and
penetrate all materials such as fireclay and alundum ; a
new form of very dense alundum, recently introduced, has
proved the most satisfactory material.

Acheson ^ has more recently introduced a soft form of
graphite for lubricating purposes. Soft graphite can be
prepared by raising the silica content of the anthracite to
65 parts of coal with 35 parts of sand. The mixture sur-
rounding the carbon-starting resistor is itself surrounded
by a mixture of carbon and sand having a still higher resist-
ance (1 part of coal to 2 parts of sand).

Soft graphite mixed with grease, oil or water is on the
market as lubricants under the names of Gredag, Oildag,
and Aquadag respectively.

For the preparation of electrodes, petroleum coke is
crushed and calcined to expel the volatile matter, then
ground in a pulverizer and mixed with pitch with a limited
quantity of petroleum, in steam- jacketed kettles. The
plastic material is pressed hot into the shape required,
usually under considerable pressures, cut into lengths,
covered with sand, and baked in a gas-fired furnace. When
graphite carbons are required 3 per cent, of oxide of iron is
added to the original coke ; the carbons are built up in the
graphitizing furnace arranged transversely to the current-
flow and packed in granulated coke for treatment. The
addition of small quantities of ammonia and gallotannic acid
is said to improve the nature of the product.

The energy consumption per kilogram of anthracite
converted into graphite can be calculated as follows : 49
Taking the mean specific heat of graphite between 20° C.
and 2200 C. as 0*45, the energy required to heat up 1 kilo-
gram of graphite will be 990 calories (0*45 X2200). The heat
evolved during transformation of the carbon into graphite
fe 236 calories per kgm., hence the total heat required is
990—236=754 cals. or o # 88 kw. hour per kgm. In actual
practice from 3 to 3*3 kw. hours are required per kgm.

158 INDUSTRIAL ELECTROMETALLURGY

Processes of electrical graphitization have been applied to
various grades of coals, and even to dried peat, but have not
proved technically successful.

Phosphorus.

Phosphorus is being produced in increasing quantities
by electrothermal methods. The process consists essentially
of smelting a mixture of bone ash or the minerals apatite,
wavelite, and rock phosphates with carbon and silica to
obtain a liquid calcium or aluminium silicate slag and
phosphorus vapour diluted with carbon monoxide, according
to the equations —

Ca 8 (P0 4 )2+3Si0 2 +5C=3CaSi03+5CO+2P
2MP0 4 +3Si0 2 +3C=M 2 (Si0 3 )3+5CO+2P

The chief difficulties associated with the electrical
production of phosphorus are those associated with its
condensation and the ease with which the phosphorus
vapour will penetrate through porous materials, even
through the furnace walls.

The first satisfactory furnaces, designed by Readman
and Parker, were operated on the resistance system, the
electrodes being disposed horizontally near the base in a
firebrick cylinder with a domed roof. The charge is con-
tinuously fed in through the roof by means of a screw
conveyer so as to exclude air, and the slag is drawn off by
intermittent tapping every three or four hours. Reduction
is said to commence at 1150 C. 60 and to be completed at
1460 C. The phosphorus is condensed in copper vessels
under water.

With ores containing but small quantities of iron, 80-90
per cent, recovery is obtained in this kind of furnace.

In later designs of furnace, such as those of Irvine,
Machalske (Anglo-American Chemical Co.) and I^andis
(American Phosphorus Co.), certain improvements have been
incorporated, eg. phosphorus and slag resisting furnace
liners made of vitrified brick set in an asbestos sodium

ELECTROTHERMAL PROCESSES 159

silicate mortar. Horizontal carbon electrodes have been
eliminated, and either one or more pendent electrodes substi-
tuted. The furnaces operate either on the arc or resistance
system, more frequently the former, the arc being formed
either between the electrodes themselves or between the
electrodes and an annular carbon ring set in the furnace
walls.

The earlier pattern furnaces had an output capacity of
80 kgm. of phosphorus per day ; the later ones are said to
be capable of producing the same amount in one hour.

According to S. Richards 61 the energy consumption for
the smaller furnaces was about 11*5 kw. hours per kgm.
phosphorus. In more modern and larger units this has been
reduced to 5 kw. hours per kgm.

Arsenic.

The electrothermal production of arsenic is being de-
veloped by the Arsenical Ore Reduction Co., applying the
Westman process to the ore deposits in Ontario.

The ore consists chiefly of mispickel, FeS 2 .FeAs 2 , a
thioarsenide of iron. On heating in a reducing atmosphere
a matte of ferrous sulphide is obtained containing any gold
or silver present in the ore. The arsenic is volatilized and
is condensed on the colder parts of the furnace.

In Westman's process the ore is heated by alternating
current between cast-iron electrodes in a furnace capable of
dealing with 90 kgm. of ore per hour. The ferrous sulphide
matte is tapped off from time to time whilst the arsenic is
removed from the furnace by a current of nitrogen gas.
The furnace space and a set of condensers forms a closed
system with a gas blower ; at the commencement of a run
air is circulated round the system and the oxygen removed
by combustion of some of the arsenic in the furnace. During
the period of volatilization of the arsenic, condensation in
the external condensers takes place. f

According to Hering, a metric ton of ore requires some
1000 kw. hours for treatment.

160 INDUSTRIAL ELECTROMETALLURGY

Carbon Disulphide.

All the carbon disulphide used in the various industries
in the United States, exceeding 2000 tons per year, is pro-
duced in E. Taylor's resistance furnaces at Penn. Yann., N.Y.
The electrical preparation of the sulphide is a great advance
over the ordinary thermal method, both as regards cost of pro-
duction, purity and absence of danger to the workmen. The
furnace (p. 161} consists of a double-walled cylinder containing
packed carbon at the base which serves as a resistor. Dense
carbon is not appreciably attacked by sulphur vapour. Fresh
carbon is fed in at the base from time to time through the
hoppers A, A, of which there are four. Raw sulphur can be
fed in through four similar hoppers at the top of the column,
B, B, and runs down the annular space between the double
walls of the column. By this means it arrives at the re-
action chamber at the same temperature at which the carbon
disulphide is formed, and serves as a heat interchanger to
cool the liberated vapours. The furnaces are each 41 feet
high, 16 feet in diameter, and built of iron, and they require
a current of 4000 amperes at 40 to 60 volts, transmitted
through four electrodes, D, D, each 25 sq. dcm. in cross-
section and 1 '2 metres long, at right angles to one another,
and situate in the base of the furnace. Still larger furnaces
are stated to be contemplated.

The molten sulphur (m.p. 115 C.) flows to the base of
the furnace, where it slowly vaporizes (b.p. 444*5° C), passing
up through the heated carbon which is maintained at from
800-1000 C. to a layer of charcoal in the tower. The
formation of carbon disulphide according to the equation

C-f-S2 ==: CS2

- •

is complete at a bright red heat.

Charcoal containing less than 3 per cent, ash is used, being
fed in through the hopper F situated at the top of the furnace.

Each furnace will yield approximately 1000 metric tons
of CS 2 before it is necessary to dismantle and clean out the
ash. The output from each furnace is about 7500 kgm.

ELECTROTHERMAL PROCESSES

161

per day, representing an output of about 12 kgm. CSj per
kw. hour. If we assume that the gases leave the furnace
at 200 C, we can calculate the theoretical energy consump-
tion necessary from the following data. The heat of forma-
tion —

LJ C+2S=CS 2

>G5»

-Carbon disulphide furnace. Perm. Yann., N.Y.

is 19,000 calories. To vaporize the CS 2 72 calories are required
per kgm., and to heat the vapour up to 200 C. we need
200 x 0-24=48 calories per kgm. The total amount of
energy necessary is therefore 250+72+48=370 calories,
equal to 0-45 kw. hour per kgm. or 2'2 kgm. per kw. hour.
The furnace thus shows an energy efficiency of 55 per cent.

REFERENCES TO SECTION IV.

1 "Met. Calculations," p. 8t>:

• Ind. Chetn., 1917, 7, 873.

■ " Iron Smelting at Trollhatten," Eng. and Min. Journal Feb., 1914.

* Mining Mag., Oct., 1910.

■ "The Electrometallurgy of Zinc," Trans. Amer. Electrochem. Soc,

1907. 18, p. 117.

i6s INDUSTRIAL ELECTROMETALLURGY

• " The Electric Furnace," p. 324.

7 Trans. Atner. Electrochem. Soc, 25, p. 176; 191 5.

• Trans. Amer. Electrochem. Soc, 24, 1913.

9 Petersen, Trans. Amer. Electrochem. Soc, 24, 1913.
18 U.S. Pat. 814050, 1904.
11 Proc Soc. des Ing. Civils de France, 1903.
18 Chem. Zeit., p. 416; 1913.
18 Met. <&» Chem. Eng., 1912, p. 281.
14 Trans. Amer. Electrochem. Soc, p. 191 ; 1914.
11 U.S. Pat. 814810, 1905.
10 Trans. Amer. Electrochem. Soc, 8, 171 ; 1905.

17 " Recent advances in the construction of electric furnaces for the

production of pig iron, steel and zinc." Ottawa, 1910.

18 Chem. Eng., 191 3, p. 380.

li Met. 6* Chem. Eng., 8, p. 209 ; 1910.

10 U.S. Pat. 768054.

81 Mining Journ., p. 909, 191 3.

11 Chem. Zeit., 86, 1192; 1912.

>s Kremakoff, Zeit. Anorg. Chem., 84, 333 ; 1901.
14 Zeit. Anorg. Chem., 84, 1 ; 1909.
18 Trans. Amer. Electrochem. Soc, 20, 191 x, p. 315.
88 Met. and Chem. Eng., 8, 277 ; 1910.

87 C.R., 116, 3549 ; 1893.

18 Elektrometallurgie, 3rd Auf. p. 519 ; 1903.

88 Bull. Technologique, 1904.

80 Metal. Chem. Eng., 9, 453; 19".

81 Trans. Amer. Electrochem. Soc, 18, 1910, p. 205.
88 Hutton, Trans. Chem. Soc, 1908, p. 1483.

88 Electrochem. Ind., 5, 239.
84 C.R., 122, 1270.
88 Metallurgie, 3549; 1900.
84 U.S. Pat. 835052 of 1906.

87 J. S.C.I. , p. 1016; 1907.

88 Stahl u. Eisen, 28, 356 ; 1905.

89 Trans. Amer. Electrochem. Soc, 24, 191 3, p. 186.

40 E. K. Rideal, " The Lighting Industry/'

41 CJl., 116, 122 ; 1897.
41 I.S.C., 98, 1483 ; 1908.

48 Chem. Zeit., p. 215 ; 1914.

44 Electrochem. and Met. Ind., 7, p. 142.

48 " The Electric Furnace," p. 281.

44 Electrochem. and Met. Ind., 7, 1909, p. 86.

47 H. Bragg, " X-Rays and Crystal Structure."

48 Electrochem. Ind., 4, pp. 343, 502 ; 1906.

48 Allmand, " Applied Electrochemistry," p. 444.

60 Hempel and Muller, Zeit. Angew. Chem., 18, 632 ; 1905,

81 Electrochem. Ind., 1, 17; 1902.

ELECTROTHERMAL PROCESSES 163

BIBLIOGRAPHY TO SECTION IV.

"The Electric Furnace," Stansfield.

"Der Elektrische Ofen," Bronn.

" Kunstlicher Graphit," F. A. Fitzgerald. 1904.

"Die Metallurgie des Zinns," H. Henniche«

Section V.-CARBORUNDUM AND THE
OXYSILICIDES OF CARBON

The reactions occurring between carbon and silica at the
high temperatures of the electric furnace are very varied,
and have led to the commercial production of many industri-
ally important compounds. The chief of these is silicon
carbide, named " carborundum " by E. Acheson, the dis-
coverer of the compound in 1891, who at the time was
under the impression that the material contained crystalline
alumina (corundum).

Carborundum is used in large quantities as an abrasive,
as bits for rock drills and the multitude of other uses that
a crystalline substance as hard as diamond can be put to.
Among the other important uses of the substance may be
mentioned its application as a deoxidant in the preparation of
steel and as an infusible liner for coal and coke fired furnaces.

Carborundum is prepared in a resistance furnace follow-
ing the general construction adopted by E. Acheson for the
production of graphite (see p. 156). Each furnace is about
10 metres long, 5 metres high and 3 metres broad, built up
of brickwork and containing the usual carbon core, about
1 metre in diameter, of f-inch crushed coke. A charge con-
sisting of an intimate mixture of the following composition : —

Sand . . . . . . 52*2 to 54*4 per cent.

Coke 35-4 to 35-1

Sawdust . . . . 10 6 to 7*0

Salt r8 to 3*5

is loosely packed round the core.

It will be noted that a slight excess of coke is used above
the proportions corresponding to the equation —

Si0 2 +3C=SiC+2CO

CARBORUNDUM 165

The function of the salt is to remove the impurities in the
coke and sand such as iron by the formation of volatile
chlorides. By the addition of sawdust the porosity of the
charge is maintained, thus allowing the carbon monoxide
formed during the reaction to escape. Each furnace con-
sumes about 2000 kw. At the commencement the
resistance is high, necessitating an applied E.M.F. of 200 to
250 volts, falling to 75 volts at the end of the run. During
the conversion the carbon monoxide liberated burns
between the joints in the brickwork at the sides and the top
of the furnace.

The temperature range within which the formation of
carborundum is possible is a very limited one, lying between
I 550° C. and 2200 C, and it is only by careful control of this
factor that successful preparation of carborundum is possible.

On dismantling a carborundum furnace a great variety
of products are obtained, formed by the interaction of the
constituents.

The exact mechanism by which the various compounds
are produced is by no means clear, but their line of demarca-
tion around the core are usually quite well defined. It is
found that the carbon core, now completely graphitized,
is surrounded by a zone of crystallized carborundum, SiC ;
then by a layer of carborundum powder; then a ring of
siloxicon, "fire sand," Si 2 C 2 0, mixed with silicon monoxide ;
and finally a skin of fritted silica.

According to Iyampen and Tucker, 1 Gillet, 2 and
Saunders, 3 siloxicon commences to be formed at 1500 C. to
1550 C, presumably according to the equations —

(1) Si0 2 +C-»SiO+CO

(2) 2SiO+3C-»Si 2 C 2 0+CO

whilst at 1820 C. silicon carbide formation commences,
being completed at 1920 C. At 2220 C, according to
these authors, dissociation of the carbide commences and is
completed at 2240 C. —

SiC^tSi+C

Although the explanation of the production of siloxicon

166 INDUSTRIAL ELECTROMETALLURGY

(Si 2 C 2 0) from silica and carbon is complete, investigators
are not in agreement as to the method of formation of
carborundum. It may be produced by the reduction of
siloxicon by carbon or silicon at a higher temperature —

(i) Si 2 C 2 0+C^2SiC+CO
(2) Si0 2 +C^SiO+CO
SiO+C^Si+CO
Si+Si 2 C 2 0^2SiC+SiO

or produced by interaction of silicon and carbon vapour,
the silicon being formed according to either of the following
equations : —

. {a) From siloxicon, SiO+C^Si+CO

(b) From carbide and silica, Si0 2 +2SiC^3Si+2CO

J . Richards 4 has calculated the probable vapour pres-
sure of carbon at various temperatures, with the following
results : —

Vapour pressure
Temperature. mm. Hg.

1820 C 0001

1920 C.
i960 C.

2000° C.

2060 c.

2100° C.

2155 c.

2215 c.
2255 C.

0005

001
0*02

0-03

005

O'lO
0'20

030

He considers that this small vapour pressure is quite enough
to account for the growth of carborundum crystals from
the interaction of silicon and carbon in the vapour form.
In the presence of an excess of silicon vapour the dissociation
of the silicon carbide formed would of course be depressed,
thus permitting of slightly higher working temperatures.

Tone 6 believes that the formation of the carbide is
brought about by the interaction of carbon monoxide and
silicon vapour according to the equation —

3Si+2CO^Si0 2 +2SiC

99
99

CARBORUNDUM 167

By the direct interaction of carbon and silica we can
write the equations for the production of silicon carbide and
siloxicon as follows : —

(1) Si0 2 +3C=SiC+2CO

(2) 2Si0 2 +5C=Si 2 C 2 0+3CO

In view of the extreme rareness of molecular reactions of
such a high order it is extremely probable that the direct
formation of these compounds does not take place, but they
are the result of a series of simpler reactions such as those
outlined above.

With Allmand 6 we may calculate the energy necessary
for the production of carborundum as follows : —

The heats of formation of silica, silicon carbide and carbon
monoxide are respectively —

Si+0 2 =Si0 2 +i8o,ooo calories.
Si+C=SiC+ 2,000
C+0=CO+ 29,200

Hence the production of 1 kilomol (40*3 kgm.) of carbo-
rundum at room temperature requires 180,000+2000—
(2 X 29,200) =»i 19,000 calories. We can further assume that
the carborundum is heated to 2100 C, whilst the liberated
CO on passing through the cold surrounding charge leaves
the effective part of the charge at 1400 C. The mean
molecular specific heats of carborundum and carbon monox-
ide between o° and 2100 C. and o° and 1400 C. respectively
are 11*3 and 7*1.

The heat required can be summarized as follows : —

To forming carborundum =119,600 calories.

To heating up SiC, 113 X 2100 = 23,700
To heating up CO, 2 X71 Xi4<>0= 19,900

Total =163,200

_ i63,20QX4'i9
~~ 40-3x3600
=47 kw. hours per
kgm. or 4700 kw.
hours per metric
ton.

168 INDUSTRIAL ELECTROMETALLURGY

In technical working about 8360 kw. hours per metric
ton of carborundum are consumed, and only 50 per cent, of
the charge is converted.

The product is manufactured by the Carborundum Co.
at Niagara (10,000 kw.), and a smaller plant at Dusseldorf,
Germany, is controlled by the same company.

Sn,UNDUM, SlI^FRAX.

Owing to the resistant properties of carborundum, the
substance being unattacked by oxygen even up to very high
temperatures or by acids, attempts have been made to pre-
pare moulded articles of the material in a compact form. It
was found that moulded carbon articles could be converted
into a semi-metallic state by exposure to silicon vapour
(" silidizing "). The resulting product, termed silundum,
retains its original shape and possesses all the properties of
carborundum. It is electrically conducting and can be used
for resistance material for temperatures up to 1200 C.

Carbon articles are packed in a charge of the composition
required for the production of carborundum in a heating
furnace of the usual resistance type. The carbon is con-
verted into silundum, beginning at the outside and con-
tinuing to a depth depending on the period of heating.
Fitzgerald has attempted to prepare articles of carborundum
by moulding crystallized carborundum into the desired
form and subsequently recrystallizing the mass in an electric
.furnace. Tone adopted "fire sand" or amorphous carbo-
rundum to the same purpose, using water glass or glue as a
temporary binder.

He 7 also investigated the action of silicon vapour on
carbon at various temperatures, and showed that between
1550 and 1820 C. the carbon was converted into amorphous
silicon carbide between 1820 and 2220 C. into the crystal-
line variety. He found evidence of the existence of solid
solutions of silicon in silicon carbide when the penetra-
tion was most effective, i.e. if silica in excess of the quantity
required for the preparation of carborundum be added to

CARBORUNDUM 169

the charge in which the articles are packed. The modifica-
tion of siltindum produced under these conditions he terms
" silfrax." The maximum penetration of silicon vapour
into pure carbon is 0*5 inch. " Silit," as prepared by the
Siemens Co. in Germany, appears to have been identical
with silundum.

Monax.

Potter, 8 who investigated the various products formed in
the carborundum furnace, isolated crude silicon monoxide
as a reddish-brown powder, to which he gave the name
" monax." It has a limited field of usefulness as a reducing
agent, a heat insulator (it has an apparent density of only
0*04), and a polisher of fair abrasive power. .It has also
been suggested to use it as a pigment and in the prepara-
tion of printers' inks.

SlLOXICON AND FlBROX.

We have already referred to the preparation of siloxicon
or silicon oxycarbide in one of the outer zones of the car-
borundum furnace. The oxycarbides of silicon were first
isolated by Cohen in 1881, 9 who obtained various different
compounds on heating silicon in a stream of carbon dioxide.
In 1903, Acheson 10 developed the preparation of " siloxicon "
for technical purposes, and not merely as a by-product in
the manufacture of carborundum.

" Siloxicon " apparently includes a series of compounds
of the general formula Si,C*0, where x lies between 1 and 7,
and averaging 2 when large samples are taken. It is an
amorphous powder highly refractory and indifferent to
most acids ; forms a suitable lining to furnace walls, but is
more easily oxidized than carborundum, a superficial silica
glaze being produced according to the equation —

2Si 2 C 2 +70 2 =2Si0 2 +4C0 2

It can be moulded and baked to form vessels of various
shapes. At high temperatures in an inert atmosphere it
decomposes into carbide and silicon as follows : —

Si 2 C 2 0=SiC+Si+CO

170 INDUSTRIAL ELECTROMETALLURGY

Tone has expressed the view that siloxicon may consist
essentially of a solid solution of silica in silicon carbide,
since on treatment with hydrofluoric acid it is freed from
silica and silicon, and a residue of amorphous silicon carbide
remains behind.

Acheson's method of preparation is carried out at the
International Graphite Co.'s works at Niagara. A carbo-
rundum furnace is used, although occasionally modified by
the introduction of multiple carbon cores instead of a single
one. The charge consists of a one-third coke and two-
thirds silica with the usual admixture of a little sawdust
and salt. As has already been noted, the formation of
siloxicons occurs at a lower temperature than that necessary
for carborundum.

Both F. Tone and E. Weintraub have succeeded in
preparing an interesting modification of siloxicon. The
material termed " Fibrox " by Weintraub consists of fine
threads of siloxicon (0*3 to 0'6/i diameter), frequently
several inches long. It serves as a remarkably efficient heat
insulator owing to the fineness of the fibres ; at the same
time, in common with carborundum, it is a good electrical
conductor. Its apparent density is said to lie between
00025 and 0*0030 (2 J to 3 gms. per litre). It can be formed
by allowing carbon monoxide slowly to diffuse into a vessel
containing silicon vapour at a temperature just below that
required for the formation of carborundum.

REFERENCES TO SECTION V.

1 /. Amer. Chem. Soc, 28, 850; 1906.

* /. Phys. Chem., 15, 213; 1911.

8 Trans. Amer. Electrochem. Soc, 21, 438; 1912.

4 Trans. Amer. Electrochem. Soc, 26, p. 194; 1914.

5 Trans. Amer. Electrochem. Soc, 26, 1914, p. 181.

6 " Applied Electrochem.," p. 491.

7 Trans. Amer, Electrochem. Soc, 24, p. 181 ; 1904.

8 Trans. Amer. Electrochem. Soc, 12, 223; 1907.

• CM., 92, p. 1508; 1881.
10 U.S. Pat. 722793.

CARBORUNDUM 171

BIBLIOGRAPHY TO SECTION V.

" Carborundum/' F. A. J. Fitzgerald. 1904.
"The Electric Furnace," Stansfield.
■'Applied Electrochemistry," A. Allmand.

Section VI.— THE CARBIDES AND BORIDES

Our knowledge of these products of the electrical furnace
is due chiefly to the work of Moissan, who isolated the first
carbide, that of calcium in a pure state, in December, 1892,
by reduction of lime with carbon. The carbides as prepared
in the electric furnace are all dark metallic-looking solids
with a crystalline fracture. Most of them react with water
to give off hydrocarbons with reformation of the oxide,

e -g- - CaC 2 +2H a O=Ca(OH) 2 +C 2 H 2 .

The products of decomposition are shown in the following
list : —

Carbide. Product.

Fe 8 C

Cr 4 C and Cr 3 C 2

Mo 2 C

W 2 C

SiC

TiC

Cs 2 C 2
Na 2 C 2
K 2 C 2
Rb 2 C 2
14 2 C 2
CaC 2
SrC 2
&42»C 2

BC 2
A1A

Mn 3 C

CeC 2

LaC 2

PrC 2

NiC 2

SnC 2

YC 2

ThC 2

U 2 C 2

\ No action with water.

C 2 H.

• ••

CH 4

CH 4 and H 2

CH 4 , H 2 , C 2 H 2 and traces of
other volatile hydrocarbons.

Volatile and liquid hydrocarbons.

THE CARBIDES AND BORIDES 173

The carbides most important industrially are those of
calcium and silicon, the former being used for the pro-
duction of acetylene and as an intermediary in the cyana-
mide industry as well as a reducing agent in electrothermal
work, the other, the preparation of which has already been
discussed, as an abrasive and in small quantities as a reducing
agent.

Calcium Carbide.

Calcium carbide is produced in electric arc furnaces
according to the following reaction : —

(1) CaO+3C$CaC 2 +CO

The reaction is reversible, and elevation of the temperature
shifts the equilibrium over in favour of the formation of the
carbide. Simultaneously with this reaction other side re-
actions take place in the furnace, viz. the formation of
calcium from the oxide and the carbide according to the

equations —
H (2) CaO+C$Ca+CO

(3) CaC 2 $Ca+2C
and the dissociation of the carbon monoxide —

2C0^2C+0 2

Several attempts have been made to determine the
pressure of CO required for equilibrium at various tempe-
ratures.

According to Allmand l the measurements of Rudolphi 2
and Thompson 3 are most correct ; the mean values found
were —

Temperature. P.CO in mm. Hg.

1575 c. . .

1675 c. . .
1775 c. . .
1875 c. . .
1975 c. . .

50
253

107-0

4050

13300

The reaction between lime and carbon begins at about
1500 C, and fusion of the carbide commences at 1800 C., 4
but in actual practice the resulting carbide is nearly always
heated to 2000 ° C. 6 At this temperature a pressure of

174 INDUSTRIAL ELECTROMETALLURGY

nearly two atmospheres of CO would be necessary to re-
convert the carbide back into lime and carbon. If the tempe-
rature be raised too high the carbide is said to be " burnt,"
dissociation of the carbide into graphite and calcium vapour
taking place according to reaction (3), the resulting calcium
acting with the carbon monoxide present to reform lime and
carbon dust, which are carried off in the gas stream.

There have been several calculations made on the theo-
retical energy expenditure necessary for the production
of one metric ton of pure calcium carbide from lime and
carbon, varying from 1523 kw. hours to 3837 kw. hours
(Gin). Several investigators have given no details as to
the temperature at which the various products (CO and
CaC 2 ) are supposed to leave the furnace, and there is still
some uncertainty in the values of the specific heats of lime
and carbon at high temperatures, as well as the heat of
formation of calcium carbide.

We will adopt the following values as the basis of calcula-
tions : —

Atomic specific heat of carbon = 4 , 26+ooo72T
Molecular specific heat of lime=n , 4+o , ooiT
Heat of formation of CaO =145,000 calories

CO = 29,200

CaC 2 = 3,900

per gram
molecule

If we assume that the carbide and carbon monoxide leave
the furnace at 2000 C, and are not used to heat up the
incoming charge, we may calculate the energy required as
follows : —

Heat ( ra * se J m °l CaO from o° to 200c C = 24,800

supplied to 1 decomposer mol. CaO into Ca and O 2 =i45,ooo

p ^ (raise 3 mols of C from o° to 2000 C.= 30,000

Heat liberated! of 1 mol CO = 29,200 199,800

by formation (of 1 mol CaC 2 = 3,900

33**00

Net energy required=i66,700 calories or

= 3,040 kw. hours per metric
ton,

THE CARBIDES AND BORIDES 175

a figure closely approximating to that which Haber obtained
(3100 kw. hours) by a somewhat different method, in which
he assumed that the value of Q in the reaction —

CaO+3C+Q=CaC 2 +CO

was equal to 121,000 — 3*3T, a value determined by Thomp-
son {loc. cit.). Commercial calcium carbide averages 85 per
cent, purity, and the energy consumption per ton of 85 per
cent, carbide varies between 3500 and 5960 kw. hours,
depending on the construction of the furnace and the system
of operation.

The carbide furnace was originally designed for heating
by means of an electric arc, but, as in most other electro-
thermal processes, the tendency to make use of " resistance "
heating has led to very considerable modifications of the
types of furnace employed and incidentally to an increased
thermal efficiency. It is, however, doubtful whether any
carbide furnace operates purely on the resistance principle,
since owing to the presence of both calcium and carbon
vapour in the hot charge, the E.M.F. necessary for striking
an arc can be reduced to 8 volts. It appears more probable
that the best furnaces act as smothered arc furnaces, in
which small arcs are continually made and broken by the
movements of the chaige. Both alternating and direct
currents have been used ; the most uniform product is
obtained with the former, but in badly designed furnaces
there is frequently a considerable loss of energy due to self-
induction.

The charge consists of a mixture of carbon and crushed
limestone in the theoretical proportions. If no loss were
maintained 1440 kgm. of the charge should produce 1000
kgm. of carbide ; in practice from 1700 to 3000 kgm.
of charge have been found necessary for this output. The
two important factors in the charge which influence the
yield and the energy consumption are : (1) Size and uni-
formity of composition, and (2) Presence of impurities.
It has been found that a finely comminuted charge is inad-
visable. The evolution of carbon monoxide during the

176 INDUSTRIAL ELECTROMETALLURGY

process of formation of the carbide causes honeycomb
channels to be formed in the interior of the mass, frequently
glazed on their interior surfaces ; this results in the loss of
a considerable amount of heat due to the ease with which the
hot gas can escape from the charge without heating it.
When the channels become numerous a mass of overlying
charge may subside, causing a large fluctuation in the furnace
load and at the same time liberating a sudden burst of gas
which carries with it a portion of the finely powdered
charge. In practice the most uniform results are obtained
by using lime crushed to i inch and the coal to J-inch or
^ inch.

The lime for carbide manufacture should be thoroughly
burnt and free from moisture. The carbon in the form of
coke, anthracite, or charcoal should have as low an ash
content as possible. Coke containing more than 10 per cent,
ash can only be used with great difficulty, and less than 5 per
cent, ash is desirable. Anthracites containing 3 per cent,
ash, or less, are usually employed. Small quantities of
carbonized wood and sawdust, the by-product of wood
distillation, find their way into the carbide industry and
give the best results.

The usual impurities are magnesia and alumina, which
assist in the formation of thick crusts in small furnaces and
cause the molten carbide in the larger furnaces to beeome
viscous and not so easily tapped ; small quantities of
arsenic, phosphorus and sulphur present in the coal or in
the lime as calcium phosphate or sulphate are reduced to
phosphides, arsenides and sulphides in the furnace.

When treated with water impure carbide will liberate
phosphine, arsine and hydrogen sulphide, all objectionable
on account of their toxic properties. In addition, impure
phosphine is spontaneously inflammable. The other coal
ash constituents, silica, oxide of iron and the alkalis, do not
sensibly affect the operation of the furnace.

Various alternative schemes have been proposed to
ensure the preparation of pure carbide from impure materials
so as not to limit the manufacturer to the purchase of pure

THE CARBIDES AND BORIDES 177

anthracite and lime. Rathenau suggested the addition of
iron to the melt to remove the silica as ferrosilicon, which
may be drawn off from the bottom of the furnace below the
carbide. Hewes adds a quantity of limestone and about
2 per cent, of manganese dioxide to the charge. The carbon
dioxide liberated from the carbonate serves to carry im-
purities such as tlie calcium sulphides and phosphides to
the surface crust, whilst the manganese carbide formed
lowers the melting point of the calcium carbide, permitting
it to be easily tapped.

The addition of crude hydrocarbon oil to the lime instead
of heating up a mixture of carbon and lime is said 6 to give
a loose non-coherent, non-hygroscopic carbide.

Calcium carbide furnaces of three distinct types are in
use at the present time —

(1) The " block " or " ingot " type.

(2) Tapping furnaces.

(3) Continuous furnaces.

(1) Ingot Furnaces. — The earlier forms of carbide
furnaces were all of the " ingot " type, such as the Willson,
Bullier and Horry furnaces. In these furnaces a charge
of lime and coke is fed in, and when a sufficient amount of
carbide is formed the furnace is removed and allowed to cool.
The contents are then broken up and the fused carbide
separated from the half-formed and non-converted material ;
the latter is returned to the furnace with the next charge.
The earlier Willson furnace (Fig. 19, A) had one basal and one
pendent electrode which was continually raised as the block
of molten carbide increased in size. Owing to the loss of
energy occasioned by transmission of the current through
the partly solidified mass, the improved Willson (Fig. 19, B)
was introduced having two pendent electrodes.

As has already been observed, only part of the charge
is converted, the unconverted and semi-fused materials
acting as a protecting liner for the iron furnace walls.
Several of such furnaces are generally run together, each
taking a charge of about 1400 kgm. and an energy con-
sumption of 200 to 250 kw. at 50 to 70 volts. About
1,. 12

178 INDUSTRIAL ELECTROMETALLURGY

800 kgm. of carbide is formed from this charge after a
13-hour run. The energy consumption of 6000 kw. hours
per metric ton of 85 per cent, carbide has been reduced to
4500 kw. hours in the later designs of the Willson furnace,
but the labour cost of breaking up and sorting the ingot is
always high.

In the Horry furnace of the Union Carbide Co., Niagara
(Fig. 19, C), a successful attempt has been made to apply
the ingot system to a semi-continuous operation. A wide
horizontal spindle carries two rings about 25 metres in
diameter and 1 metre apart. By means of movable plates

acKd
car bide

A " -0 ' c

Fig. 19. — Calcium carbide furnaces. "Ingot" type.

the space between the plates can be converted into a chamber
of rectangular cross-section in which the reduction of the
carbide takes place. Two electrodes are mounted in a
hopper supplying the charge at one point in the ring, the
outside plates being then removed for the purpose. The
spindle is slowly rotated so as to remove the molten carbide
from the base of the hopper as rapidly as it is formed, thus
allowing a fresh charge to accumulate above the old one.

After a complete rotation the cover plates are removed
and the ring of solidified carbide broken up. A complete
revolution is made in 24 hours and about two tons of carbide
are produced. With a load of 375 kw, per furnace 7 a

THE CARBIDES AND BORIDES 179

production of 1 metric ton per 4500 kw. hours is thus
obtained. According to Stansfield, 8 the energy consumption
can be brought down to 310 kw. per furnace, thus pro-
ducing 1 metric ton for 3800 kw. hours.

(2) Tapping Furnaces. — With an increase in the furnace
size to reduce the radiation heat loss and the labour cost per
ton of material, a corresponding increase in the efficiency
of working and ease of control of temperature was obtained.
The practicability of tapping a large mass of molten carbide
impossible in the smaller furnaces owing to the formation of
crusts and the high viscosity of the impure melt was investi-
gated in many carbide works.

The Alby Carbide Co. at their Odda works use a furnace
very similar to the early Willson pattern provided with
tapping holes in the end walls. Each unit will take 1000
kw. at 50 volts and is tapped once every 45 minutes.
A considerable economy is effected both in current and in
raw material, 1500 kgm. of charge being used to produce
1 metric ton of carbide as against a theoretical charge of
1400 kgm. The energy consumption is stated to vary from
4200 to 4500 kw. hours for 1000 kgm. of carbide.

(3) Continuous Furnaces. — Several improvements have
been made in tapping furnaces so as to ensure continuity
of production, especially by Helfenstein, 9 by Memmo in
Italy, and in the Norwegian carbide works. When large
open arc furnaces are used a limit is set to the power con-
sumption by the difficulties of working and the nuisance
from fumes. The present tendency is to use large multiple-
phase current furnaces, which are totally enclosed. By this
means the radiation loss from the upper surface of the charge
is minimized and the opportunity of collecting the carbon
monoxide evolved presents itself.

Helfenstein has used a 9000-kw. 3-phase furnace,
furnished with three electrodes, one to each phase, with
satisfactory results. The practical limit to the current
consumption in an enclosed furnace is set only by the size
of the electrodes, which should not carry a current exceeding
500 amperes per sq. dcm.

180 INDUSTRIAL ELECTROMETALLURGY

Borchers suggested water cooling the furnace for the
purpose of steam generation with the waste heat ; this idea
does not seem to have received technical application. The
carbon monoxide leaving the furnace at 2000° C. can be
utilized for burning the limestone and preheating the entering
charge.

The energy liberated in the combustion of one gram-
molecule of carbon monoxide is approximately 67,700
calories, whilst the energy required to heat up 3 molecules
of carbon and one molecule of CaO in accordance with the
equation —

CaO+3C=CaC t +CO

from o" to 2000 C. is, as we have seen, only 54,800 calories.

Fig. 30. — Calcium carbide furnace. Three-phase continuous type.

It should, therefore, be possible to heat up the charge to the
reaction temperature by the combustion of the carbon
monoxide liberated during the formation of the equivalent
amount of carbide. By preheating the charge to 2000 C.
the consumption of energy required for the production of
1 metric ton of carbide would be reduced to 2100 kw. hours.
The carbon monoxide can, of course, be burnt after
parting with its heat to the incoming charge, as indicated
in the accompanying diagram. The energy thus derived

THE CARBIDES AND BORIDES 181

may be used for raising steam or burning the limestone.
In some modern furnaces an auxiliary basal electrode, D,
is fitted, through which a current can be supplied during
the process of tapping. By this means a high temperature
is maintained at the tap hole, and a steady fluid stream of
molten carbide is obtained.

Water cooling the electiode holders has further minimized
the consumption of electrode material. Sufficient evidence
is not yet to hand as to the highest electrical efficiency obtain-
able with this type of furnace, but individual experimental
runs have shown the possibility of producing a metric ton
of carbide with an energy consumption of only 3800 to
4000 kw. hours.

The preparation of borides is not accomplished on an
industrial scale. In view of the remarkable abrasive powers
of certain borides which exceed carborundum and alundum
in hardness, an outlet for a small supply of a high-grade
material might be found. Their method of preparation
is on a small scale similar to that adopted for calcium carbide.

REFERENCES TO SECTION VI.

1 " Principles of Applied Electrochemistry/* p. 420.

* Met. Chem. Eng., 8, 279 ; 1910.

3 Trans. Amer. Electrochem. Soc, 9, 158 ; 1900.

4 Hansen, Electrochem. 6* Met. Ind., 7, 1909, p. 427.

5 Borchers and Rothmund, Zeit. Elehtrochem., May, 1902.

• Wright, " Electric Furnaces and their Industrial Applications," p. 62.

7 J. Richards, Electrochem. Ind., 1, 22; 1902.

8 " The Electric Furnace," 1913, p. 303.
» Trans. Faraday Soc, 5, p. 254 ; 1910.

i82 INDUSTRIAL ELECTROMETALLURGY

BIBLIOGRAPHY TO SECTION VI.

"Principles of Applied Electrochemistry," A. Allmand.
"Electric Furnaces and their Industrial Applications/' Wright.
"The Electric Furnace," Stansfield.
"Der Electrische Ofen," Bronn.
•' Carbide of Calcium," C. Bingham.

Section VII. ^ELECTROTHERMAL
NITROGEN FIXATION BY METALS AND

METALLIC COMPOUNDS

In 1898 Sir William Crookes, in his presidential address to
the British Association, drew the attention of the scientific
world to the growing importance for a satisfactory solution
of the nitrogen problem. 1 National and international
economic and political existence all centre around the land
question, and we find that fixed nitrogen is an essential
for the production of food from the land. Approximately
four-fifths of the world's supply of nitrogenous materials
are used as fertilizers, the remaining one-fifth in the chemical
industries, chiefly as cyanides for the extraction of gold,
as nitric acid for the production of explosives, and in various
forms for the diverse branches of the organic chemistry
industry, especially the dyes.

With the natural increase in the density of the popula-
tion, a corresponding increase in intensive horticulture is
necessary, and we find that Belgium, one of the most densely
populated areas of the world, uses more nitrogenous fertil-
izers per acre than any other country, and a corresponding
increased yield per acre of foodstuffs is obtained, whilst in
the almost virgin soils of the wheat areas in America,
Canada and Siberia the application of any nitrogenous
fertilizer has not yet been found necessary. Apart from
these natural tendencies towards an increased consumption
of artificial fertilizers, the development of certain social
factors, such as the system of peasant proprietorship en-
suring a greater interest in the land, the increased scientific
education of the people leading to a more rational view as to
the needs of the soil ; and the increased power of purchase

184 INDUSTRIAL ELECTROMETALLURGY

by means of guilds and co-operative societies, all indicate
that the demand for artificial fertilizers is certain to increase
at a greater rate than it has done in the past.

The present sources of supply may be briefly enumerated
as follows : 2 —

Nitre. — Large deposits of natural sodium nitrate,
"caliche," are found in Chile. These have been worked
on an ever-extending scale since 1830, the present output
amounting to over two and a half million tons per annum.
Although these deposits appeared inexhaustible when first
worked, the increasing consumption has led to various
carefully investigated surveys of the area to determine
the probable available supplies. The reports submitted
to the Chilean Government have exhibited a diversity of
opinion, and although the somewhat alarming figure of
21 years as the maximum period of life can be rejected, it
appears that before the end of the present century the eco-
nomic exportation of Chilean nitre will no longer be possible.

Less important deposits of nitrates of sodium and
potassium are found within the British Commonwealth,
namely in India, the Sahara, Egypt and Persia, and will
probably be developed locally.

Ammonia. — The other chief source of supply of combined
nitrogen is ammonia, which is used in the form of ammonium
sulphate. Recently experiments by Rossi and others have
indicated that ammonium nitrate can also be used as a
fertilizer. It may be remarked that since ammonia can
be converted into nitric acid by combustion with air or
oxygen on the surface of suitable catalytic material or by
oxidation in aqueous solution, the production of a fertilizer
from ammonia is not necessarily dependent on the sulphuric
acid industry. Practically all the natural ammonia avail-
able is recovered from coal distillation either from gasworks
or coke ovens. Only about 20 per cent, of the fixed nitrogen
in the coal is obtained in the usual distillation process,
but a somewhat better recovery is obtained in Mond Gas
Producers (60 per cent.), whilst still smaller amounts are
recovered from shale distillation and blast furnaces.

ELECTROTHERMAL NITROGEN FIXATION 185

The world's annual coal production is over one thousand
million tons, of which Great Britain supplies over 15 per cent.
Since coal contains on the average 1 per cent, of nitrogen,
with a 20-per-cent. recovery two million tons of fixed nitrogen
in the form of ammonia would be available if at least
partial carbonization of all coal were made compulsory.
It is evident that the present available supplies from this
source alone could entirely replace the Chilean nitre if the
requisite legislation were introduced. Although the world's
coal reserve is larger than the Chilean caliche, their period
of economical working is by no means indefinite.

Other sources of ammonia are found in the by-products
of the destructive distillation of bones and the organic
residues in fermentation industries. Probable sources of
natural fixed nitrogen will possibly be found in the extended
development of the gasification of the low-grade fuels such as
peat and turf, of which extensive areas are found within
the British Empire, especially in Canada and Ireland ;
and in the more extensive application of sewage works
sludges to manurial purposes. 3 At the present time these
potential sources of natural ammonia are not being eco-
nomically developed ; the bulk of the nitrogen in the coal
is entirely lost owing to the extravagant methods of fuel
combustion employed, whilst the technical difficulties
associated with the drying of sewage sludge, peat and turf
to render it suitable for fuel have not yet been satisfactorily
solved. As a consequence, during the last ten years there
has been a considerable development in methods suitable
for the technical fixation of atmospheric nitrogen. These
may be classified as follows : —

A. The direct preparation of nitric acid by oxidation
of the atmospheric nitrogen in the electric arc, and by
combustion of gaseous fuel.

B. The preparation of synthetic ammonia from its
elements.

C. The fixation of atmospheric nitrogen by biochemical
methods.

D. The preparation of combined nitrogen from which

t86 INDUSTRIAL ELECTROMETALLURGY

ammonia can be obtained, e.g. cyanamides, cyanides and
nitrides.

A. (i) The Arc Process. — The pioneer work connected
with this method was accomplished in England by the
investigations of Cavendish 1781, Davy 1800, Rayleigh
1897, an d McDougall and Howies, who erected the first
technical plant in 1899 at Manchester. An outline of the
present industrial arc processes, the operation of which is
practically confined to Norway, although smaller plants
are in operation in Switzerland, Italy, France and Germany,
are discussed in another volume of this series. It may,
however, be pointed out that the electrical efficiency of the
process is very poor, the yields obtained in technical processes
being only from 65 to 75 gms. of nitric acid per kw. hour.
The efficiency of a normally operating furnace of the
Birkeland, Schonherr or Pauling type can be calculated
as follows : The energy required to form a gramme-molecule
of nitric oxide according to the equation N 2 +0 2 =2NO is
approximately 22,000 calories, and although 13,500 cals.
are liberated during the formation of nitrogen dioxide
according to the reaction 2NO+02=2N0 2 , this heat
evolution occurring during the last stages of cooling the
gases and absorption in water is not technically available
as recoverable energy. Since 22,000 cals. are equivalent
to 00256 kw. hour, this amount of electrical energy
must be supplied to form 30 gms. of nitric oxide or 63 gms.
of nitric acid. Taking the mean of Nernst, Jellinek and
Haber's figures for the equilibrium concentrations of nitric
oxide in ordinary air at high temperatures, the mean absolute
temperature of the air passing through the arc to give a
2 per cent. NO concentration must be in the neighbourhood
of 2500 C. It must, however, be remembered that since
equilibrium is obtained with great rapidity at these high
temperatures a considerable higher mean gas temperature
may be attained in practice, and that the low percentages
of gas actually obtained are due to the impossibility of
cooling the gases sufficiently quickly to relatively low tempe-
ratures. At low temperatures the apparent stability of a

ELECTROTHERMAL NITROGEN FIXATION 187

gas rich in NO is assured owing to the slowness with which
the reverse reaction 2NO =N 2 +0 2 proceeds. Assuming that
a 2 per cent, concentration of NO represents the true value
of the equilibrium concentration at 2500 C, the energy
necessary to heat up 14 gms. of nitrogen and 16 gms. of
oxygen to this temperature is approximately 2500 X2X

(6-8+0 0006X2500) , - . *u. 4.4.1 •

— - or 20,600 calories. The total lrrecover-

able energy consumption is therefore 42,600 calories, or

0*050 kw. hours for the production of 63 gms. of nitric

acid, representing an electrical efficiency of only — — or

5 per cent, for a technical production of 63 gms. of nitric
acid per kw. hour.

The remaining 95 per cent, of the electrical energy passes
out with the unchanged air. Partial recovery of this loss
is effected by passage of the arc gas through Babcock and
Willcox boilers for raising steam. Although no data are
available as the amount of energy so recovered, there can
be no doubt that the process is extremely wasteful in
power. Its chief merits are extreme simplicity and uni-
formity of action. The erection costs are high owing to
the large volume of gas that has to be heated, cooled and
passed through absorption towers.

The work of Haber and Koening on chilled arcs, of Rossi
on the use of arc furnaces worked under reduced pressure,
and more recently of Lowry and Strutt on the production
of an allotropic active modification of atmospheric nitrogen,
indicate that arc methods may be capable of modification
and improvement, possibly departing from the original
electrothermal process to an electronic one so as to render
it suitable for those countries where the cost of electrical
power is the governing factor in electrochemical develop-
ment.

(2) By the Combustion of Gaseous Fuel.— Chiefly
owing to the work of Haber and his pupils on the pro-
duction of oxides of nitrogen during the process of gaseous
explosion and combustion, a number of patents have been

188 INDUSTRIAL ELECTROMETALLURGY

taken out for utilizing gaseous fuel for the direct formation
of oxides of nitrogen. The only one which has been de-
veloped on a semi-technical scale is that of Hausser, in
which a mixture of coke oven gas and air or oxygen is
exploded in a bomb of large capacity. Provision is made
for a rapid cooling of the gases by water injection and rapid
release into the cooling system. The process of charging,
release, ignition and scavenging, can be made cyclic on a
modified Otto cycle.

The mean explosion temperature of 2100 C. when
operating with rich coke oven gas should yield an NO con-
centration in the resulting gases of only 05 per cent, or
about 5-56 gms. per kw. hour. The inventor has put
forth claims to obtaining over 5 per cent, of NO, due to the
induced photochemical action favouring the formation of
endothermic compounds during the period of explosion.
These claims, however, remain to be substantiated.

B. The Haber Process. — The successful technical
development of the experiments of Regnault and Ramsay
on the synthesis of ammonia from the elements nitrogen
and hydrogen was made by Haber and the Badische Anilin
u. Soda Fabrik at Oppau and Leverkusen in Germany a
few years previous to the outbreak of war. Haber first
determined the amount of ammonia in the equilibrium
mixture 3H 2 +N2$2NH 3 , under high gas pressures and at
various temperatures. It will be noticed that from the
equilibrium constant equation —

C»h, l C Ni

increase of pressure increases the equilibrium amount of
ammonia present, and furthermore an excess of hydrogen
over the stoichometric ratio H 2 : N 2 : : 3 : 1 is likewise
beneficial for high concentrations of ammonia.
From the general equation —

° g K 2 RYf 2 TV

ELECTROTHERMAL NITROGEN FIXATION 189

the dependence of K, the equilibrium constant, on the tem-
perature can be determined provided the heat of formation
of ammonia be known. Haber obtained the following con-
centrations in close agreement with the calculated values,
tinder a pressure of 200 atmospheres in an equilibrium
mixture of the three gases : —

Temperature.

Percentage NH 8

350°

117

750°

2-99

950°

1*07

At low temperatures the velocity of conversion is slow,
whilst at high temperatures the equilibrium amount of
ammonia formed is reduced. The two chief obstacles
overcome by Haber were, the construction of suitable
furnaces to withstand high pressures and the preparation
of a catalyst which would bring about the combination of
the two gases at low temperatures. Suitable catalysts
working extremely actively at very low temperatures
400 to 500 C. were found in osmium and uranium carbide,
but difficulties were encountered when technically prepared
gases were employed owing to the poisoning of the catalyst
by the small traces of impurities such as carbon monoxide,
sulphur compounds, oxygen, and the like in the gases. Cata-
lysts more robust but less active were found in electrolytic
iron and certain ferrugineous mixtures, such as iron and
molybdenum, iron and tungsten, iron and cobalt, ferro-
cyanides and specially prepared sodamide metal mixtures.

The general arrangements of the Badische plant are
fairly well known, but details as regards construction and
catalytic material employed are carefully guarded national
secrets. The hydrogen is prepared by the " Bamag "
process, in which water-gas and steam is passed over a
specially prepared iron catalyst at 550 C, when the following
reaction proceeds to equilibrium : —

H 2 +C0 +H 2 0^2H 2 +C0 2

The equilibrium amount of CO remaining in the gas at this

190 INDUSTRIAL ELECTROMETALLURGY

temperature, calculated from the equilibrium constant
K= n - C ^-^ 1 is about 2 per cent. The excess steam is

removed by condensation, the C0 2 and any sulphur now
present as H 2 S by pressure washing with water, and the CO
is converted to formate by scrubbing with hot caustic soda
under pressure. Small traces of the monoxide are removed
by a cuprous ammonium carbonate scrubber. The nitrogen
is prepared from liquid air or alternatively by so adjusting
the air and the steam blasts in the water-gas producer to
preparer a nitrogen-hydrogen mixture in one operation.
The purified gases are now compressed and passed over
palladium asbestos and calcium chloride to ensure the
absence of traces of oxygen and water-vapour.

The catalyst " bombs " were originally heated externally,
but owing to the weakening of the metal by the combined
action of the hydrogen and the high temperature, internal
electrical heating is now applied. The gas circulates through
the bomb, which contains the catalyst and a system of heat
interchange coils. Since the reaction N 2 +3H 3 ->3NH 3 is
exothermic, local over-heating of the catalyst may occur,
when the amount of heat generated automatically falls owing
to decomposition of ammonia already present. The
electrical energy necessary for maintaining the catalyst
temperature is small when a good system of heat inter-
change is installed. The ammonia formed in the gaseous
mixture is subsequently removed by liquefaction or counter
current washing with water under pressure, and the unused
gas, augmented by an additional supply from the compressor,
dried and recirculated through the bomb. Periodic
" blowing off " is necessary to eliminate inert gases such as
methane from the hydrogen, and argon present in the
nitrogen. The velocity of circulation of the gases is im-
portant for the economic production of ammonia. At very
low speed equilibrium is obtained, but the yield per litre
of catalyst space is small owing to the low velocity of the
gas-flow. At very high speeds equilibrium is not obtained,
and only a low percentage of ammonia is formed, but the

ELECTROTHERMAL NITROGEN FIXATION 191

yield per hour is higher owing to the increased velocity of
the gas-flow. The space time yield, i.e. the yield in kilo-
grammes of ammonia per litre of catalyst per hour, is
the determining factor as far as output of a unit is con-
cerned. In technical operation the space time yield may
rise as high as 15. Naturally with increased velocity of
gas-flow, the difficulties of heat regeneration and gas circula-
tion are greatly enhanced. The cost of the process is chiefly
determined by the expense entailed in the preparation
and purification of the hydrogen, the compression of the
gases and the skilled supervision necessary.

(C) Biochemical Nitrogen Fixation. — In ordinary
soil a vast number of bacteria are present, frequently rising
to more than 10 million per gramme. These include amongst
the several varieties of saprophytic bacteria both aerobic
and anaerobic, pathogenic organisms and moulds, a number
of organisms capable of fixing atmospheric nitrogen, entirely
distinct from the nitrifying bacteria of Winogradsky and
the more recently discovered denitrifying organisms which
are capable of oxidizing or reducing ammonia, nitrates and
nitrites already in the soil as such or as nitrogenous organic
substances. The first organism possessing this property
of fixing nitrogen was isolated by Benjerink, viz. Aostridium
Pastorianutn. Since this period a large number of organisms
have been shown to possess this property, such as the fungus
Aspergillus niger, Penecillium glaucum, Phoma beta and
others, and among the bacteria Azotobacter chroococcum
and agilis, and Bacillus radiciola. The conditions for
successful nitrogen fixation in soil are briefly as follows : —

1 . Presence of calcium, phosphorus and smaller quantities
of sodium and potassium.

2. Large quantities of fixed nitrogen hinder further
absorption.

3. The temperature range lies between io° and 50 C.

4. The earth should be well aerated and not contain
less than 15 per cent, of moisture.

Most of the organisms function most successfully when
growing in symbiosis with other organisms. The azobacter

192 INDUSTRIAL ELECTROMETALLURGY

grow most abundantly with certain algse whilst Hellriegal
showed that the B. radicicola was practically only associated
with leguminous plants such as peas ; in this case the plant
and bacillus exhibit alternate parasitism. Of recent years
several strains of nitrogen-fixing organisms have been
cultivated on artificial media for agricultural purposes,
and during the period of the war it has been stated that
large quantities of nitrogen-fixing yeasts have been grown
in Germany for supplying pigs and other animals with their
requisite organic nitrogen. In the inoculation of soils the
choice of the bacillus employed should be determined by
the nature of the contemplated crop, and the activity of
the organism controlled before use, since this deteriorates
when grown for several generations in artificial media.
A good growth is ensured by the addition of a little bacterial
pabulum to the soil, such as grape sugar and peptone.
Up to the present time these methods have scarcely proved
commercially successful.

D. Electrothermal Fixation of Nitrogen by Metals
and Metallic Compounds. — Elementary nitrogen has been
fixed by purely thermal processes in several forms on an
industrial scale. Amongst the more important may be
mentioned the nitrides, the cyanides and the cyanamides.

All these compounds can be converted into ammonia
by treatment with water or steam under a few atmospheres
pressure according to the following equations : —

X 3 N +3H 2 =NH 3 +3XOH
X 2 NNC+3H 2 0=X 2 C0 8 +2NH 3 (under 6 atmospheres

pressure)
2X 2 NNC+2H 2 0=(H 2 NNC) 2 +2H 2 in water
XCN +2H 2 =XCOOH +NH 3
XCN+2H 2 0=XOH+NH 3 +CO (at 500 C.)

The Nitrides. — The technical fixation of nitrogen has
been accomplished by Serpek in the Savoy, and although
several difficulties prevented the process from being eco-
nomically successful, yet the advantages of such a process
of nitrogen fixation is so great that a reinvestigation of
the problem would probably prove remunerative.

ELECTROTHERMAL NITROGEN FIXATION 193

Serpek's early experiments (1 906-1 907) were devoted to
the preparation of aluminium nitride by passing nitrogen
over aluminium carbide at a red heat, when according to
Caro 4 the following reactions take place : —

A1 4 C 3 $4A1+ 3 C

4A1+2N 2 ^4A1N

The dissociation temperature of the nitride is higher than
that of the carbide, and consequently the nitride is actually
formed through the intermediary preparation of aluminium.
From 1907-1910 Serpek was engaged in the construction
of apparatus suitable for carrying out the following reaction
on a technical scale : —

A1 2 3 +3C +N 2 =2 A1N +3CO

The temperature at which the nitride decomposes relatively
quickly is 2120 C, and consequently the reaction must
take place below this temperature. Serpek states the
absorption of nitrogen commences at 1100 C. ; at 1500 C.
the absorption is fairly rapid, whilst from 1700 to 1850 C.
the reaction is a comparatively violent one. We have
already noted that alumina purified by the electrothermal
method is not readily soluble in the electrolyte used for the
preparation of aluminium. A similar observation was made
by Serpek in connection with the formation of nitride.
Bauxite is more readily converted to nitride than alumina,
and absorption commences at a much lower temperature.
Tucker and Read 6 confirmed these results and came to
the conclusion that low-temperature fixation, desirable owing
to the relative ease with which the nitride is again dis-
sociated, can be brought about by suitable catalysts usually
present in French bauxite. Serpek successfully operated
two different types of furnace in the Savoy. In the first
a rotary kiln was employed having transverse carbon
resistors heating a charge of bauxite and coke by radiation.
The second type of furnace consisted of a hollow vertical
shaft containing an axial carbon resistor. The annular
space is charged with a mixture of bauxite and coke, and

1.. 13

194 INDUSTRIAL ELECTROMETALLURGY

a current of nitrogen passed through the charge at a tempe-
rature of 1600 to 1700 C. The liberated carbon monoxide
is burnt to heat up the incoming charge. Pure alumina
can be recovered from the nitride by decomposition with
superheated steam or weak alkali —

MN+3H 2 0=A1(0H) 3 +NH 3

Attempts to use the alumina after dehydration did not
prove successful, probably owing to the absence of suitable
catalysts. Although the alumina could of course be eco-
nomically used in the preparation of aluminium, yet the
future development of a technical nitrogen fixation process
on these lines is more likely to be successful if a continuous
process could be devised so as to render the process inde-
pendent of the need of a supply of fresh bauxite for each
charge. From both Serpek and Tucker's observations,
it appears possible that a synthetic catalyst could be
prepared presumably containing other oxides, such as iron
titanium or chromium, which would permit of the suc-
cessful utilization of the alumina.

Another possible development of the process in which
the alumina is treated as a catalyst for gaseous reactions
may be imagined as follows : At a temperature of 1500 C.
methane is nearly completely dissociated into carbon and
hydrogen according to the reversible equation —

CH 4 ^C+2H 2

whilst carbon monoxide may be converted into methane by
hydrogenation over a nickel catalyst (provided the CO,H 2
mixture contain less than 10 per cent. CO) —

CO+3H 2 =H 2 0+CH 4

when comparatively low temperatures are employed, i.e.
under 380 C. At higher temperatures carbon is deposited
according to the equation —

CO+H 2 ^C+H 2

Either of these reactions obviously permit of the prepara-
tion of carbon in a finely divided form. It would therefore

ELECTROTHERMAL NITROGEN FIXATION 195

seem possible to alternate a steam blow and a nitrogen
carbon blow or a nitrogen, hydrocarbon blow through a shaft
containing alumina especially sensitized with suitable cata-
lysts maintained at the requisite temperatures. It has been
stated that the presence of hydrogen in the gases facilitates
the fixation of nitrogen at low temperatures, and these
conditions would be maintained in the above imaginary
process. The theoretical power consumption for the Serpek
process is small, since the reaction —

A1 2 3 +3C +N 2 =2 A1N +3CO

requires only 213,000 calories per kgm. mol or per metric
ton of fixed nitrogen calculated as ammonia 7300 kw.
hours.

It will be noted that 1*5 mols of carbon monoxide are
formed per molecule of the nitride, whilst in the case of
calcium carbide manufacture only one mol CO per mol
CaC 2 is obtained. The gaseous utilizable energy by com-
bustion of the CO is consequently far greater.

If we assume that a reacting temperature of 1500 C.
can be utilized by the choice of specially prepared alumina,
the energy necessary to heat up one kilomol of alumina,
3 kilomols of carbon and 1 of nitrogen to the reacting
temperature is about 40,000 calories. By the combustion of
1 "5 kilomols of carbon monoxide 102,000 calories are obtain-
able. It follows that theoretically 62,000 cals. are avail-
able for the process of nitrogen absorption, reducing the
amount of energy to be supplied from 213,000 to 151,000
calories, or per metric ton of fixed nitrogen calculated as
ammonia from 7300 kw. hours to 5170 kw. hours.

Various patents have been taken out for the preparation
of ammonia from other nitrides besides aluminium, especially
magnesium, silicon, boron and titanium. There is no
evidence that any of them have passed the experimental
laboratory scale, and the principles involved are the same
as those used by Serpek which have already been dis-
cussed.

The Cyanides. — Since the market value of fixed

196 INDUSTRIAL ELECTROMETALLURGY

nitrogen in the form of cyanide is greatly in excess of its
actual value as a fertilizer, 6 and that at present the most
successful synthetic cyanide process, viz. the Castner,
utilizes expensive raw materials, sodium, ammonia and
charcoal, according to the equations —

2NH3 +2Na =2NaNH 2 +H 2
NaNH 2 +C=NaCN+H 2

a great number of processes have been proposed for the
production of cyanides utilizing atmospheric nitrogen.
Up to the present time these processes have not been able
to compete with the existing ones in which some form of
combined nitrogen is used as a starting-point. It is evident
that many difficulties would have to be overcome before
a cyanide nitrogen fixation process could be developed
not only sufficiently economical in operation to compete
with the cyanide processes already extant, but providing
a sufficient margin in working costs so as to permit of the
cyanide so produced to be sold at fertilizer prices after
having been converted into some transportable form of
ammoniacal nitrogen. During the last few years when the
nitrogen shortage has stimulated research, attention has been
redirected to these synthetic cyanide processes, and there
are prospects that one of these newer modifications, viz.
the sodium cyanide process, developed by Bucher, may
finally supplant both the synthetic ammonia and the calcium
cyanamide processes. In this section only those methods in
which electric heating has been used or suggested will be
considered ; this will naturally exclude a large number,
yet those included will be representative and appear to be
those which would be most feasible for application on a
technical scale.

Cyanogen and Hydrocyanic Acid. — Berthelot first
indicated the formation of hydrocyanic acid by passing
electric sparks through a mixture of acetylene and nitrogen.
The union of acetylene and nitrogen to form hydrocyanic
acid proceeds more smoothly when a diluent such as hydrogen
is used.

ELECTROTHERMAL NITROGEN FIXATION 197

Gruszkrewicz obtained his best yield with a gas mixture
of the following composition : —

5 per cent. C 2 H 2
5 per cent. N 2
90 per cent. H 2

Better results were obtained when water gas was employed.
A 0*3 per cent, conversion was obtained in an hour with a
gas mixture containing

54-5 per cent. CO
25 per cent. N 2
20*5 per cent. H 2

The use of the electric arc for the production of hydro-
cyanic acid has recently been reinvestigated. Lepinski
claims a . 19 per cent, conversion by passing a gas of the
following composition through an arc : —

70 per cent. N 2
20 per cent. CH 4
10 per cent. H 2

whilst the Neuhausen Aluminium Co. are said to employ
a gas of the mixture —

5-10 per cent. CH 4
66-81 per cent. H 2
12-24 per cent. N 2

The reaction proceeds smoothly above 1800 C, but the
methane content should not exceed 10 per cent, owing to
the formation of large quantities of soot. The hydro-
cyanic acid appears to be formed when the thermal decom-
position of the hydrocarbon occurs —

CH 4 =2H 2 +C

This reaction only occurs above 1300 C, whilst the decompo-
sition of the hydrocyanic acid itself also proceeds rapidly
at high temperatures. Successful development of these
processes similar to those employed in the ordinary oxide
of nitrogen arc would seem to be indicated where a rapid
chilling of the gases after heating is obtained.

198 INDUSTRIAL ELECTROMETALLURGY

No data are available as to the conversion efficiencies of
any of these processes, but it is evident that if a 03 per cent,
yield obtained by Gruszkrewicz can be converted into a
19 per cent, yield as claimed by Lepinski by simple transition
from a spark discharge to a high temperature arc discharge
when utilizing crude producer gas or producer gas en-
riched with some hydrocarbon, a cyclic process of heating,
heat interchanging and scrubbing the gases with an alkali
would certainly prove worthy of investigation.

The Alkaline Earth Cyanides. — Readmann 7 first
suggested the use of electrical heating for the fixation of
nitrogen by means of a mixture of alkaline earth oxide or
carbonate and carbon. As suitable mixture he suggested
the following : —

BaC0 3 . . . . . . . . 50 kgm.

Charcoal . . . . . . . . 10 kgm.

The intimate mixture of these substances is introduced into
a coke-lined crucible and is heated to a high temperature
by the passage of an electric current introduced by means
of carbon electrodes inserted in the sides of the crucible.
Deoxygenated air or water gas is passed through the mixture.
When the latter is used the denitrified residue may be used
as fuel. The absorption was said to proceed smoothly
according to the equation —

BaO+3C+N 2 =Ba(CN) 2 +CO

Part of the cyanide so formed flows out through a lateral
opening situate in the bottom of the crucible and part
volatilized with the gases, from which it can be recovered
by absorption. The optimum temperature of absorption
was 1400 C. The above process was slightly modified
by Swan and Kendall, in which titanium, molybdenum,
chromium or manganese was previously added to the
charcoal alkaline earth mixture before absorption of nitrogen.
It is stated that with these catalysts the formation of
cyanide will commence at a dull red heat, thus avoiding
the high temperatures necessary when no catalyst is present.

ELECTROTHERMAL NITROGEN FIXATION 199

Mehner suggested the f ollowing ingenious process : Fused
barium cyanide is electrolyzed between granulated carbon
electrodes in an atmosphere of nitrogen gas. Cyanogen
is liberated at the anode and can be absorbed in water or
caustic alkali; the barium set free at the cathode reacts
at the temperature of the melt with the granulated carbon,
preferably charcoal, and the nitrogen to reform barium
cyanide, which is thus continuously reformed.

The Alkali Cyanides. — The observations of Mond
confirmed the earlier experiments of Possoz and Boissiere,
who erected the first nitrogen fixation factories in the world
at Grenall and Newcastle in 1843, of Newton, Swindel, and
Margueritte and Sourdeval, in that both the alkalis and
alkaline earths readily reacted with carbon and nitrogen
to form cyanides at high temperatures. Barium reacted
most easily, and many unsuccessful attempts have been made
to modify Readmann's process so as to render it com-
mercially feasible. Amongst the alkalis potassium reacted
more easily than sodium, and the formation of potassium
cyanide was noticed by Dawes and Clarke in the Clyde
blast furnaces as early as 1835 and 1837. Thompson in
1839 fast drew attention to the catalytic effect of iron, and
the catalytic effects of other metals, such as manganese
and chromium, were noted by Margueritte and Sourdeval
in i860, and by Swan and Kendall in 1895. Bucher 8 has
recently investigated the process, and claims to obtain rapid
absorption of the nitrogen in producer gas at a temperature
of 900 C, by briquettes of sodium carbonate, coke and iron
as a catalyst. Decomposition of sodium cyanide by means
of superheated steam is complete at 6oo° C, according to
the following equations : —

Na 2 C0 3 +4C +N 2 =2NaCN +3CO
NaCN +2H 2 =HCOONa +NH 3
2NaCN+4H 2 0=Na 2 C0 3 +2NH 3 +CO+H 2

To heat up 1 kilomol of sodium carbonate, 4 kilomols of
carbon and one of nitrogen to 900°C. requires 43,000 Calories;
the reaction itself is slightly endothermic, 140 Calories being

200 INDUSTRIAL ELECTROMETALLURGY

absorbed, and the energy available from the combustion of
3 kilomols of CO is over 200,000 calories. It is therefore
evident that the reaction as a whole can be considered as
a strongly exothermic one, and in practice should be capable
of continuous production without the supply of any ex-
traneous energy.

The carbon consumption per metric ton of ammonia
produced would be equal to 1300 kgms. It will be noted
that the working temperatures of the cyanide processes are
much lower than those necessary for the formation of
nitrides or of the carbides, necessary intermediaries for the
preparation of cyanamides.

The Cyanamides. — Frank and Caro, in 1895, investi-
gated the preparation of cyanides through the intermediary
formation of the alkaline earth carbides, and were the first
to study the conditions favourable to the formation of
cyanamides. According to the authors the reaction usually
expressed by the equation —

CaO +3C +N 2 =Ca (CN) 2 +CO

really takes place in several stages —

1. CaO+C^lCa+CO

2. Ca+2C^CaC 2
3(a). CaC 2 +N 2 ==CaCN 2 +C

(6). CaC 2 +N 2 ^Ca(CN) 2

The following conditions were found favourable to the
formation of calcium cyanamide : —

(1) The nitrogen should be in excess of the theoretical
amount required.

(2) A porous carbide is desirable.

(3) Relatively high temperatures should be employed ;
for calcium cyanamide 1100 C. appears to be the optimum.

With the overproduction of calcium carbide which took
place at the beginning of the century the possibility of
utilizing these observations of Frank and Caro for the
fixation of atmospheric nitrogen immediately presented
themselves, and at the present time the cyanamide industry
is a large one, the annual production being nearly half a

ELECTROTHERMAL NITROGEN FIXATION 201

million tons exclusive of the increased production of the
Central Empires during the period of the war. Leblanc °
investigated the formation of cyanamides of calcium and
barium, and came to the conclusion that the relative amount
of cyanamide and cyanide formed were dependent on the
degree of dissociation of the carbide —

1. CaC 2 ^CaC+C
CaC+N 2 =CaCN 2

2. CaC 2 +N 2 ^Ca(CN) 2

Absorption was found to commence at 700 C. in the
case of both barium and calcium carbide, but the barium
product always contained large quantities of cyanide.
Polzenius and Carlson carried out a series of experiments
on the use of suitable addition agents to facilitate the
absorption of nitrogen. The chloride and fluoride of cal-
cium were found to give the best results.

Cyanamide of calcium is now manufactured by two
processes, the intermittent and continuous. The con-
tinuous process is more economical than the earlier inter-
mittent ones and produces a somewhat higher grade product.

In the intermittent process at work at Odda in Norway
the carbide is crushed and ground to a fine powder and
packed in small sheet-iron drums, each of 300 to 500 kgms.
capacity and more recently of 1 to 2 metric tons capacity.
The drums, which are lined with refractory bricks, are heated
internally by carbon resistors, each drum taking some
20 amperes at 70 volts. Pure nitrogen, which should not
contain more than 0*4 per cent, of oxygen prepared by
fractionating liquid air by the Linde or Claude processes
or by the passage of air over hot copper, is passed in when
the temperature has risen to 8oo° C, and when absorption
commences the current is reduced and finally turned off,
the temperature being maintained by the reaction, which is
exothermic. The absorption is complete after about 30
hours' passage of the gas, during which period the tempe-
rature is maintained at from 800 ° to 1000 C. In practice
it is found that 1 metric ton of cyanamide is produced

202 INDUSTRIAL ELECTROMETALLURGY

from 078 to o-8o ton of carbide. The contents of the
drums are allowed to cool, and after crushing and packing
are placed on the market as " Nitrolim," containing some
15 to 20 per cent, of fixed nitrogen.

The electrical energy expenditure required for heating
the carbide is about o*i kw. hour per kilogram or 500 kw.
hours per metric ton of nitrogen fixed. According to
recent estimates, the cost of production of one metric ton
of nitrogen by fractionation of liquid air should not exceed

Comparison of Nitrogen Fixation Processes.— The
various estimates for comparing the costs of fixing nitrogen
by the alternative processes already described really offer
no guide to the solution of the fundamental problem of the
conservation and most economical utilization of the natural
resources of the world. It is evident that if large supplies
of nitre existed in some nearly inaccessible region, the cost
of that nitre delivered to the consumer would be too great
to permit of their economical development. In the modern
practice, since manual labour can be almost entirely replaced
by mechanical work, the international solution of the
problem will be decided by one consideration only, viz.
the energetic efficiency of the process taken as a whole,
i.e. the power and material costs involved in taking nitrogen
from the air and fertilizing the soil in the chosen areas with
it. The national problem is somewhat different, the choice
of a method is not solely determined by the efficiency of
the process in terms of the energy required to fix a given
amount of nitrogen and the transportation costs to the
consumer, but the relative costs of the different forms of
energy available for that nation, erection and running costs
become the deciding factors. In the various processes
alluded to two sources of energy are usually required,
viz. electrical and carbon in the form of coke ; the relative
costs of the energy at the factory in these two forms would
be thedeciding factors between two equally efficient processes
or alternatively between two different proposed factory sites.

In the following table the approximate consumption

ELECTROTHERMAL NITROGEN FIXATION 203

of energy and coke required to fix 1 metric ton of nitrogen
by the different processes are given : —

Kw. hours per metric O

3ke kgm. per

Other source

Process

ton nitrogen.

metric ton N 2 .

of energy.

Arc

60,000-65,000

Nil

Hatisser . .

Nil

Gas, 30,000
cu. m. Cal.
value 4300
per cu. m.
=3 150,000
kw. hrs. or
100,000
kgm. coal.

Haber . .

For gas com-

For prep.

•

pression

1,200

ofH 2

•

For N 2 pre-
paration

700

For heating ca-

Water gas

talyst

600

production 2,400

For circulation

"Bamag":

H 2

of gases

200

products

on 1,600

2,700

4,000 kgm.

Serpek . .

For N 2 pre-

paration

700

1,300

For reduction

and azotising

9,000
9>700

•

Bucher . .

For N 2 pre-

paration

700

1,710

Cyanamide

For production

of the carbide

15,000

1,400

Forazotizing . .

500

For grinding . .

700

For production

of the nitrogen

700

16,900

It is interesting to note that the Haber process con-
sumes a relatively large amount of coke for the production
of the hydrogen, and that any technical development of

204 INDUSTRIAL ELECTROMETALLURGY

either the Bucher or modified Serpek (see p. 1Q2) processes
would be serious rivals to the Haber or cyanamide. 10

REFERENCES TO SECTION VII.

1 " The Wheat Problem." London. 3rd edit., 1917.

* See also " The Alkali Industry/' Partington, this series.
8 S. Rideal, " Sewage Purification."

* Zeit. Angew. Chem., 28, 2412.

* Trans. Amor. Electrochem. Soc, 24, p. 64 ; 1912.

6 Over nine times the value in 1914.

7 French Patent of 1895.

8 /. Ind. Eng. Chem., 9, 233 : 1917.

9 Zeit. Elektrochem., 17, 20, 194.

10 Reviewed in the " Alkali Industry," J. R. Partington, this series.

BIBLIOGRAPHY TO SECTION VII.

"The Cyanide Industry," R. Robine & M. Lenglen. Wiley & Sons.
1906.

"The Fixation of Atmospheric Nitrogen," J. Knox. 19 14.

"Fabrication Electrochemique de l'Acide Nitrique et des composes
Nitres," Escard. 1901.

" Die Technische Ausnutzung der Atmospharischen Sticks toff," Donath
u. Frenzel. 1907*

" Utilization of Atmospheric Nitrogen," T. Norton. Washington, 191 2,

" Coal Tar and Ammonia," Lunge.

' ' Technologic der Cyan verbindungen, ' ' Bertelomann. 1 906.

Section VIII.— IRON AND THE FERRO-
ALLOYS

Electrolytic Iron. — The preparation of electrolytic iron
has during the last few years been fairly established as a
practical industrial process. There is an increasing demand
for pure iron of 99*95 per cent, to 99*97 per cent, purity
for the manufacture of electric machinery, where a highly
inductive iron with a low hysteresis is desired, and for coat-
ing metals, e.g. boiler tanks, as a protection from corrosion. 1

Iron is similar to nickel in its electrochemical behaviour.
Since its electrolytic potential in n. ferrous ion solution is
E A =s=4-o # 34 volt, and the hydrogen overpotential at the
metal surface is very low, in alkaline solution ?j=oo8 volt,
the deposition of hydrogen can scarcely be avoided. As
in the case of nickel, the evolution of hydrogen is assisted
by the cathodic passivity in the discharge of the ferrous ion.

To obtain the maximum efficiency it is therefore
desirable that the ratio C F€ /Ch should be kept as high as
possible, provided always that basic salts are not formed ;
that high current densities and high temperatures should
be employed to raise the hydrogen overpotential as far as
possible, to increase the velocity of discharge of the ferrous
ion as well as to decrease the solubility of hydrogen in the
deposited metal. 2 The electrolytic deposition of iron has
been developed by the firms of Mercke of Darmstadt and
Iyangbein and Pfanhauser at Leipzig, by Cowper Coles
in this country, and by Burgess in America. Both chloride
and sulphate electrolytes are stated to give good results.

Chloride Electrolytes. — E. Mercke 3 employs a solu-
tion containing equal quantities of ferrous chloride and
water (100 gms. of moist crystals in 75 gms. of water)

206 INDUSTRIAL ELECTROMETALLURGY

warmed to 65 C. as electrolyte. Good deposits are obtained
with a current density of 3 to 5 amps, per sq. dcm., provided
that the electrolyte is circulated. Wrought iron anodes
are employed, and the E.M.F. is about o*6o volt. I^ess
than 01 per cent, of hydrogen is occluded in the deposited
metal. Langbein and Pfanhauser likewise use a chloride
electrolyte containing calcium chloride in addition (700 gms.
CaCl 2 , 600 gms. FeCl 2 to 1 litre of water), at a higher tempe-
rature (90 C), and with a higher current density from 15
to 20 amperes per sq. dcm.

Sulphate Electrolytes. — Burgess and Hambuechen 4
have deposited iron of 99*97 per cent, purity with a very
high current efficiency (over 90 per cent.) from a ferrous
ammonium sulphate solution containing 40 gms. iron per
litre with the addition of ammonium chloride, at a tempe-
rature as low as 30 C. With iron anodes of Swedish bar
or American ingot and a current density of 06 to 1 ampere
per sq. dcm. an applied E.M.F. of 1 volt was found necessary
to overcome the passivity at both anode and cathode.
Storey 6 gives the following analysis of an iron produced
under these conditions : —

Fe, 99963 per cent. ; H 2 , 0083 per cent. ; C, 0013 per cent.
P, 0020 per cent. ; S, 0001 per cent. ; Si, 0*003 P er cent.

O. P. Watts a and H. I4 find that a mixed sulphate
and chloride electrolyte containing 150 gms. crystallized
ferrous sulphate (7H2O), and 75 gms. ferrous chloride (4H2O),
per litre is better than either sulphate or chloride electrolyte
alone. As suitable addition agents they advised 6 gms.
ammonium oxalate or 0*5 gm. hexaminetetramine (formal-
dehyde ammonia) per litre of electrolyte.

Other electrolytes, such as complex tartrates, citrates
and oxalates, have been suggested from time to time.
Classen's 7 electrolyte, containing ferrous ammonium sul-
phate, an equal weight of oxalic acid, and 7 times its weight
of ammonium oxalate, is the only one from which carbon-
free iron can be deposited and this only when the author's
procedure be followed in detail.

IRON AND THE FERRO-ALLOYS 207

Cowper Coles has patented the use of iron salts of several
aromatic acids for the deposition of electrolytic iron.

The I&bctrothermai, Production of Iron.

A. Ore Smelting. — Pig iron is generally produced in
a blast furnace by smelting oxide of iron ores with coke
or charcoal and limestone. The fuel is burnt at the base
of the furnace.

The ordinary blast furnace can be conveniently divided
into five zones : —

(1) The top zone, in which the entering charge is heated
by the ascending hot gases. In this zone part of the carbon
monoxide produced in the lower zones is oxidized to carbon
dioxide —

2CO+0 2 ^2C0 2

the heat liberated assisting to warm the incoming charge ;
to drive off the water in the ore, and convert the calcium
carbonate into oxide.

(2) In this zone the ferric oxide is reduced to ferrous
oxide by the carbon monoxide —

Fe 2 3 +2CO =2FeO +C0 2 +CO

(3) In the third zone, where actual reduction to the
metal takes place, the temperature is sufficiently high to
bring about the reduction —

FeO+C=Fe+CO

The metal melted in the third zone, together with the
gangue of the ore, chiefly silicious, and containing the ash
of the fuel fluxed with the lime to form a fusible slag, flow
to the base of the third zone, where they separate into two
immiscible layers, the slag on the top of the molten metal.
Both slag and pig are tapped off at intervals, fresh charge
being admitted at the top of the furnace shaft.

The gases leaving the furnace are still hot (from 300 C.
to 8oo° C), and combustion of the carbon monoxide into
carbon dioxide is not complete, the ratio CO to C0 2 being

208 INDUSTRIAL ELECTROMETALLURGY

about 2:1. The waste gas is generally used for steam
raising.

In the ideal furnace the temperature is so controlled
that only the oxide of iron is reduced and not the other
impurities, such as silica, manganese and the phosphates.
In actual practice the phosphorus, half the manganese, and
a small quantity of silicon and sulphur are retained in the
iron.

The following analyses of iron ores represent those used
in actual practice : —

Limonite.

Hematite.

Magnetite

Fe 2 3 .

• 73-840

88-500

—

Fe s 4 .

—

78*400

MnO .

• 0567

0*470

0700

Si0 8 .

I^O

2690

8-650

A1 2 3 .

2-152

3*470

7330

CaO

0-590

0870

2*100

MgO .

—

I 030

0-070

0*078

0-055

0-124

0*093

0-008

Power Consumption. — In the production of pig iron
the quantity of coke or charcoal used as fuel is about equal
to the amount of pig iron produced ; charcoal of course
being preferable, owing to the absence of impurities such as
sulphur and silica present in the coke. In the electric
furnace the carbon is used only for reduction of the oxide,
and not for heating the charge; the fuel consumption is
therefore reduced to about one-third of the fuel required
for blast furnace operation. To heat the charge electrical
energy has to be supplied, and in good modern practice
2000 kw. hours will produce one ton of pig. If we assume
that the working costs of the two systems are the same and
that the same quality of pig is produced, electric smelting
will become cheaper than blast furnace pig when 2000 kw.
hours of electrical energy cost less than 0*66 ton of high-
grade coke or charcoal.

The thermochemical data necessary for calculating

IRON AND THE FERRO-ALLOYS 209

more accurately the theoretical efficiency of the ore-smelting
process can be broadly summarized as follows 8 : —

If we take an ore containing 90 per cent. Fe 2 O s , 2 per
cent, of water, and 8 per cent, of other impurities, and find
by experiment that an easily tappable slag is obtained by
the addition of 12 per cent, of limestone to each ton of ore
in the charge, we can calculate the necessary energy expendi-
ture for this reduction. It necessarily follows that if lower-
grade ores are used, more flux will be required and a corre-
sponding increase in wasted energy for slag production
will result.

The heat of production of one metric ton of iron accord-
ing to the equations

Fe 2 3 +CO =2FeO +C0 2
2FeO +2C =2Fe +2CO

is found equal to —

(Fe 2 ,Q 8 )-(C,0 8 )-(C,0) calorfes per k . lomol

Since — (Fe 2 ,0 3 ) =201,000 calories

(C,0 2 )= 97,200
(C,0) = 29,200

the energy required per metric ton of metal is 565,250
calories.

To heat the iron so produced up to the melting-point,
to melt it and to bring it to the tapping temperature, a
further 350,000 calories are required.

For each ton of molten pig produced (from 16 tons of
ore) 200 kgms. of limestone will be required.

For calcining 200 kgms. of limestone approximately
85,000 calories will be required, and 240 kgms. of slag will
be produced (112 kgms. CaO+128 kgms. other impurities
in the ore). For reducing the impurities, heating and
melting the slag to bring it up to tapping temperature,
600x240=144,000 calories will be required. The total
heat required for slag formation is therefore 239,000 calories.

If we assume that the gases leave the furnace at 500 C.

and contain carbon monoxide and dioxide in the ratio of

i*. \ ::>l 14

. . - •

210 INDUSTRIAL ELECTROMETALLURGY

2:1, the energy carried off by the gases can be calculated
as follows : —

For every kilomol of metal 1 kilomol of carbon monoxide
and 0-5 kilomol of carbon dioxide are produced according
to the equations given above. Taking the molecular
specific heat of CO as 6*9 and of C0 2 as 100, the heat lost

in the gas is 500 X69 calories in the CO and calories

in the C0 2 , or per metric ton of pig produced 53,000 calories
in the CO and 38,500 in the C0 2 , a total of 91,500 calories.
To this must be added the energy available in the sub-
sequent combustion of the carbon monoxide present in
the gas, viz. 68,000 calories per kilomol, or 1,046,000 calories
per metric ton of pig, and the heat lost in the evaporation
of the 2 per cent, of water and superheating the steam to

500 C, i.e. - — — x 637 x 1000 +- — — X 400 X 0*48 x 1000

100 °' 100 ^

=30,000 calories, making a total of 1,167,500 calories.

Hence, for the production of one ton of pig from iron ore

of the composition indicated above, the distribution of the

energy required is as follows : —

Total energy required for the production of

the metal . . . . . . . . . . 915,250 cals.

Total energy required for the production of

the slag . . . . . . . . . . 239,000

Total energy lost in the gases . . . . 1,167,500

Total .. 2,321,750

For every kilomol of metal produced 1 kilomol of carbon
is theoretically required, or per metric ton of pig 214 kilos
of carbon.

For calculating the total energy required for the produc-
tion of 1 metric ton of pig, we have the following figures : —

Energy required for metal production . . 915,250 cals.
Energy required for slag production . . 239,000

Energy lost as heat in the gases . . . . 121,000

Total . . 1,275,250
equivalent to 150Q kw. hours for a very high grade ore.

IRON AND THE FERRO-ALLOYS 211

If fuel be used for heating the furnace, air must be
injected for supplying the oxygen; the resulting gas will
therefore contain large quantities of nitrogen, and owing
to its greater velocity will leave the furnace at a higher
temperature, viz. 900 ° C. If we assume that fractional
combustion of the carbon proceeds to the same C0 2 : CO
ratio, viz. 1:2, as in the electric furnace, we shall require
sufficient air to complete the following reaction : —

3C+8N 2 +20 2 =8N 2 +2CO+C0 2

The heat liberated by the combustion of 3 kilomols of carbon
to CO and C0 2 in the above ratio is 155,600 calories, and the
heat absorbed by the gases leaving at 900 C, taking the
molecular specific heat of nitrogen as 7, is 72,000 calories,
giving a net heat of combustion for 3 kilomols of carbon of
83,000 calories. To produce 1,275,250 calories we would
therefore require 550 kilos of carbon ; for reduction of the
iron oxide we have seen that 214 kilos of carbon are re-
quired, making a total of 764 kilos.

In this case the total energy lost in the gases owing to
the increased production of carbon monoxide is much
greater, viz. 4,390,000 calories, than when electrical heating
is used.

As to how much of the available heat in the effluent
gases can be feasibly utilized for preheating or power pro-
duction wide variations are found in practice. It will be
noted that in the electrothermal process 50 per cent, of the
total energy supplied electrically and as fuel is thus lost.
In the usual thermal process, over four times the quantity
requisite for the actual smelting operation is lost. The
gaseous products from the electric furnace consist of practi-
cally a pure CO and C0 2 mixture, whilst the blast furnace
gas is diluted down with a large amount of nitrogen, the
gas in the first case consisting of two-thirds combustible
CO, with one-third of diluent, while in the second only
two - elevenths combustible CO with nine - elevenths of
diluent.

Furnaces. — The earliest successful furnaces employed

212 INDUSTRIAL ELECTROMETALLURGY

for the production of pig iron were those of Keller
and Heroult, originally used for the production of ferro-
alloys.

The Keller furnace consists of two pendent electrodes
in vertical shafts communicating by a passage, CC. The
charge is fed in through hoppers placed round the electrodes,
and provision is made for drawing off the escaping gases.
Two tapping holes are provided, one for the metal, B, and
the other for the slag, A. A basic dolomite lining bound
with tar was found the most convenient lining. In Dr.
Haanel's report to the Canadian Government (1904), he
details an experimental furnace at Livet (France), where

(A) H^<«ir (

Fig, si. — Furnaces for thu productj

n and ferro-alloys.

over 30 tons of ore were smelted during his visit. Using
a good quality of hematite ore (48-1 per cent, iron), he was
able to produce either white or grey cast iron at will, ex-
ceedingly low in sulphur, with an energy expenditure of
2200 kw. hours per ton, using 360 kgms. of 91 per cent,
carbon coke and 0"i7 kgm. of carbon electrode.

In the Heroult furnace but one pendent electrode is
employed. A carbon base plate continued as a liner to
a little above the level of the slag serves as the other elec-
trode. Slag and metal are removed by the tapping holes
A and B. During the period of smelting fresh charge is
fed in round the electrode, which is gradually raised. Dr.
Haanel reported favourably on this furnace in operation
at Sault Ste Marie (U.S.A.), producing pig iron from such

>» >t a a

ff f> ft i>

IRON AND THE FERRO-ALLOYS 213

ores as hematite, magnetite, pyrrhotite and titaniferous iron
ores with the following energy expenditures : —

Hematite . . 1800 power consumption per ton pig in kw. hrs.
Magnetite . . 1900
Pyrrhotite . . 2570
Titaniferous v
iron ore 5 35

Using low-grade charcoal as a reducing agent, 500
kgms. per metric ton of pig produced were required. The
electrode consumption varied from 14 to 16 kgms. per ton
of pig.

Modifications of the Keller and Heroult furnaces have
been introduced by Harmet, Haanel, Stansfield, and others,
but these do not include any radical change from the original
designs.

Shaft Furnaces. — A distinct advance in the design of
furnaces for iron ore reduction was introduced by Lyon
in California and Gronwall, Lindblad and Stalhane in
Sweden, who realized that the construction of large units
in which the shafts were practically obstructed by the
electrodes was impracticable.

Lyon's earliest types of furnace were modifications of
Heroult's in which a series of pendent electrodes passing
through the roof of a smelting chamber alternated with
charging hoppers for supplying fresh charge. Preheating
of the charge was attempted by burning the carbon monoxide
evolved from the smelting charge round the hoppers.
Although difficulties in operation, such as the blocking of
the hoppers by the heated ore, led to a fresh series of experi-
ments on the lines indicated by the Swedish engineers, yet
this type of furnace is in operation in California by the
Nobel Electric Steel Co., and by Helfenstein. It is stated 9
that the power consumption per metric ton of pig varies
from 2200 to 3000 kw. hours, using a magnetite ore con-
taining some 70 per cent, of iron and 400 kgms. of charcoal
per ton of pig produced.

In Sweden, preliminary experiments by Messrs. Gronwall

214 INDUSTRIAL ELECTROMETALLURGY

Tjndblad and Stalhane at Domnarfvet led to the erection
of a series of large shaft furnaces at TrolMtten. These
operated in a highly satisfactory manner, and furnaces of
this design were subsequently erected at Hagfors in Sweden
and at Hardanger and Arendal in Norway in varying sizes,
from 3000 kw. to 7500 kw. per furnace.

The furnace shaft is 15 metres high, whilst the hearth
has a maximum diameter of 4
* metres and is 2 metres high,
constricted to a diameter of
1 metre 25 cms. where it
enters the shaft. Both shaft,
roof and hearth are constructed
of a steel shell lined with fire-
brick, with an inner lining of
magnesite bound with tar.

Four or six equally spaced
electrodes 60 cms. in diameter
serve to conduct the current to
the hearth.

Three-phase current is used,
transformed at the furnace from
10,000 to 50 to 90 volts pressure
and 12,000 to 20,000 amperes
on each phase. The thermal
efficiency of these furnaces is
stated to be nearly 80 per cent.,
and an output of nearly one ton
of pig per hour can be obtained.
The fuel consumption per ton of
pig produced is from 03 to 07 of a ton with an electrode
consumption of 4-5 kgms. for reduction of a magnetite ore
(50-60 per cent. iron). Provision is made for removing the
dust and scrubbing part or all of the evolved gases, which
consist of carbon monoxide and dioxide in the ratio 2 : 1,
together with a little hydrogen and a smaller quantity of
nitrogen. The clean gas is returned by means of a blower
to the hearth through tuyeres, whilst the unscrubbed excess

Ttppint *of «

ElictToiUs

ia. 22. — Shaft furnace for

IRON AND THE FERRO-ALLOYS 215

gas is piped away and used for steam raising. A high
carbon dioxide content has a deleterious effect on the elec-
trodes owing to partial combustion of the carbon, according

to the reaction —

C+C0 2 ->2CO

The power consumption of these furnaces averages 2200 kw.
hours per metric ton of pig iron. A very high grade
material is produced analyzing some 3*5 per cent, carbon,
with sulphur and phosphorus usually less than o*oi per
cent. The silicon content varies with the nature of the
ore, but can be usually maintained at less than 0*5 per cent.
It is extremely probable that future development of ore
reduction furnaces will be on the lines of the shaft type,
with circulation of the gases. Under the present conditions
of operation, the quantity of heat abstracted from the smelt-
ing chamber and carried to the charge in the shaft by the
gas circulated is not sufficient to preheat the charge to the
same degree as in the ordinary blast furnace, and further-
more, the chilling of the smelting chamber is an obvious
disadvantage. It is evident that if air were injected
together with some of the liberated gas through tuyeres
situated just above the smelting chamber itself, the heat
of combustion of the carbon monoxide could be more use-
fully employed inside the furnace than outside for steam
raising.

B. The Production and Refining of Steei,.

Although the application of the electric furnace for the
production of pig iron has been slow in development and
is still confined to areas where the price of electric power
is very low, the electrical production and refining of steel
is already a large industry, and in time will probably entirely
supplant the open hearth and Martin processes.

Electrical furnaces are employed for several distinct
purposes : —

(a) Refining open hearth and Bessemer or acid con-
verter steel with the aid of a flux.

216 INDUSTRIAL ELECTROMETALLURGY

(b) Fusion of pure materials.

Power Consumption. — Neumann 10 has calculated the
necessary energy consumption for the production of one
metric ton of steel, using various raw materials. When the
charge is inserted cold, the energy supplied includes that
necessary to heat the charge to the melting-point, to melt
it and raise it to the tapping temperature, and to reduce
any oxide of iron present. He further assumes the pig
iron to contain y6 per cent, carbon, i*68 per cent, silicon,
i # i per cent, manganese, and 0*02 per cent, phosphorus,
and the steel produced to contain 0*96 per cent, carbon
and 0*28 per cent, silicon. The heat of oxidation of these
impurities is subtracted from energy necessary for the

refinin g- Kw. hrs. per

Materials used. metric ton of steel.

Cold pig iron and flux 500

Iyiquid pig and flux . . . . . . 190

670 kgms. pig

210 kgms. ore, cold . . . . . . 460

45 kgms. lime

285 kgms. scrap
Same charge molten . . . . . . 230

675 kgms. pig

350 kgms. scrap, cold . . . . . . 280

Same charge molten . . . . . . 53

365 kgms. pig

650 kgms. scrap, cold . . . . . . 330

Same charge with molten pig . . . . 210

The Function of the Slag. 11 — The process of steel
refining is intimately bound up with the reactions which
occur both in the steel and in the slag. The metal and the
slag above it form two separate phases, and in the course
of purification homogeneous reactions may take place in
each phase, whilst heterogeneous reactions at the surface
of contact take place between slag and metal. The slag
functions both as a protector and as a refiner to the
underlying metal. Both sets of reactions require a high

IRON AND THE FERRO-ALLOYS 217

temperature to ensure a high diffusivity and a low viscosity.
The ease with which a high temperature is obtained is one
of the distinct advantages of the electrothermal processes.
The upper limit of the temperature is set by the boiling-point
of the metal, when mingling of the slag and metal occurs,
taking a long time to separate.

The chief impurities, phosphorus, sulphur, and silicon,
are removed by a varied series of chemical reactions, amongst
which the following are most important : —

Dephosphorization. — The removal of the three chief
impurities, phosphorus, oxygen and sulphur, usually takes
place in the order named. Phosphorus is removed by
selective oxidation at low temperatures ; at 1350 C. it can
be more easily oxidized than either silicon or carbon.
Oxidation is brought about by the ferric oxide which is
present in the slag at the period of its formation. As in
the case of sulphur, the phosphorus is distributed between
the slag and the metal in a definite ratio, consequently
when it is oxidized in the slag, more phosphorus enters
from the metal. The oxidized phosphorus is retained in
the slag if basic as calcium phosphate.

During the process of phosphorus removal, part of the
sulphur may be volatilized as sulphur dioxide.

Deoxidation. — When the removal of phosphorus is
complete and the temperature is elevated, the ferric and
ferrous oxide together with any manganese and nickel
oxides and at high temperatures the oxides of chromium,
tungsten and vanadium in the slag are reduced by the carbon
to the respective metals, which then return to the metal
phase. Silicon oxide is only reduced at very high tempe-
ratures. Any oxide of iron in the metal is continuously
absorbed by the slag owing to the disturbance of the
partition equilibrium, and is there reduced to metal.
Part of the oxide can also be reduced in the metallic phase
itself. To remove the last traces of oxide rapidly, various
reducing agents can be added to the metal and the oxide
formed slagged off. Aluminium, silicon as ferrosilicon and
calcium carbide have all been used for this purpose. It is

218 INDUSTRIAL ELECTROMETALLURGY

evident that during this period of reduction there exists
a danger of phosphorus being returned to the metal from
the slag by reduction of the phosphate. This can be
obviated by removal of the dephosphorizing slag before
reduction, or by rapidly raising the temperature during
the actual reduction period to form the endothermic calcium
phosphide, Ca 3 P 2 , which is not reabsorbed by the steel.

Both sulphur and phosphorus require basic slags to
effect their removal, but the removal of oxygen can be
effected in an acid slag using a silica brick lining.

Desulphurization. — The sulphur present in the original
iron divides itself between the slag and the molten metal
in a definite ratio in accordance with the general principles
of the partition coefficient ; since the ratio

solubility of FeS in slag
solubility of FeS in metal

increases with rising temperature, a high temperature for
sulphur removal is essential. Removal of the sulphur is
partly effected by oxidation to sulphur dioxide, during the
period of phosphorus removal, but chiefly due to reactions
taking place in the slag ; more sulphur diffusing from the
metal to re-establish equilibrium. Desulphurization is
brought about in the slag by means of silicon, carbon, lime
and carbide according to the temperature of the melt.
Silicon is most active at lower temperatures, whilst carbide
formation and desulphurization by means of the carbide
formed only occurs at very high temperatures. When lime
is used as a desulphurizing agent a large excess must be
present, since the reaction —

CaO+FeS^FeO+CaS

is a reversible one. The elimination of sulphur is practically
complete at high temperatures, when carbide is formed
owing to the removal of the ferrous oxide from the slag —

2CaO +2FeS +CaC 2 ->2CO +2CaS +2Fe

Prior to this the usual reaction —

CaO+FeS+C=CO+Fe+CaS

IRON AND THE FERRO-ALLOYS 219

takes place. At low temperatures the ferrous oxide in
the slag can only be removed by means of an added reducing
agent, such as silicon, usually added in the form of ferro-
silicon. This entails an extra expense, and may cause too
much silicon to be present in the resulting steel. The
intermediate formation of silicon sulphide —

2FeS+Si=SiS 2 +2Fe

probably also plays a part in the removal of sulphur by
added ferrosilicon.

Composition of the Slag. — Liquid steel leaves the
furnace at about 1550 C. to 1600 C, and in the furnace
itself the temperature attained lies probably between
1600 C. and 1700 C. during the last period of sulphur
removal. At these temperatures the slag must be perfectly
fluid, since unnecessary elevation of the temperature shortens
the life of the furnace lining. In normal furnace opera-
tion the softening point of slags lies between 1200 C. and
1400 C. The work on the composition and melting-points
of the various materials used for liners has largely been
accomplished by the Geophysical Laboratory at Washington,
but not so much work has been accomplished on the in-
fluence of the composition on the melting-point and viscosity
of the slags themselves.

Vogt and Doelter l2 showed that excess of lime or
silica in the slag raised the viscosity, whilst the addition
of calcium fluoride made slags more fluid. In the electric
furnace a 75 per cent, lime slag is still tappable. Recently
the Bureau of Mines, Washington, have been investigating
this problem, and a preliminary report has been given by
Feild. 18 He gives the following data of the softening-
points of various technical slags : —

SiO,.

Percent.
Al.O,. CaO.

composition.
MgO. CaS.

MnO.

Softening tem-
perature °C.

2*0

O'l

.. 1244

44
38
38

9
10

40
40

2*7
31

2-4

02
1*2
0*2

. . 1279
. . 1262
. . 1263

220 INDUSTRIAL ELECTROMETALLURGY

Per cent.

composition.

Softening tern

5iO,.

Al.O,.

CaO.

MgO.

CaS.

MnO.

peratuie C

22 .

• 1297

3*i

07 .

• 1331

0*5 •

• 1352

0-3 .

• 1342

o*6 .

• 1343

0*5 •

• 1358

2*9

03 .

• 1365

4'4

O'l .

• 1356

o\5 •

• 1383

5'9

o'5 .

• 1425

3'5

0'2 .

• 1403

5'5

0*2 .

. 1388

4'i

0*3 .

. 1410

He determined the tapping temperature of various slags
and found it to lie between 1470 and 1572 C. as deter-
mined by optical pyrometer and also by thermocouple.
The figures refer to blast furnace slags, and, as we have
noted, the temperature in the electric furnace is consider-
ably higher. The average viscosity of the slags at 1500 C.
was found to be about 301 times greater than that of water
at 20° C.

The pure silicates have the following melting-points : —

FeSi0 3
MnSi0 3
CaSi0 3
Mg 2 Si0 4

1050 C.
1150 C.
1200 C.
1400 C.

Types of Furnaces employed. — Three types of
furnaces have been employed for steel production and
refining, viz. the Arc, Induction, and Resistance furnaces,
but only the two former are in operation on a large scale.
The arc and induction furnaces have each distinct advan-
tages but at the same time have faults peculiar more to
the principle of heating than to the actual type of furnace
employed. In the induction furnace the metal is relatively
hotter than the slag, although it never actually attains the
temperatures obtained in the arc furnace.

IRON AND THE FERRO-ALLOYS 221

In addition, owing to the action of the electromagnetic
field the fluid metal is always moving, and a very intimate
slag metal contact is produced. In the arc furnace the
slag is relatively hotter than the metal. It consequently
follows that homogeneous slag reactions proceed best in
the arc furnaces, and the heterogeneous metal slag reactions
in the induction type.

Dephosphorization, which proceeds with a reasonable
velocity at relatively low temperatures, is usually not
complete in the arc furnace, but proceeds most smoothly
in the inductance. For the removal of sulphur where
high temperatures of slag and metal and a perfectly re-
ducing atmosphere are desirable, the arc offers advantages
over the inductance type. Furthermore, although the
latter requires less attention than the former, it suffers from
the additional disadvantage of possible emulsification of
the slag in the liquid metal, owing to the spin produced
by the electromagnetic field.

Arc Furnaces. — Of the more important types of arc
furnaces employed may be mentioned the Girod, the H6roult,
Keller and Stassano's.

The Girod 14 furnace is representative of the conducting
hearth furnace in which the current passes from one or
more pendent electrodes through the slag and metal to
the hearth. A combination of arc and resistance heating
is thus obtained.

A good number of furnaces of this type are at present
in operation in Europe and America, from J ton up to
12 tons capacity. Even larger sizes are in contemplation.

The lining of the furnace is usually calcined magnesia
or dolomite bound with pitch. If the temperature be
carefully controlled it is stated that nearly 100 charges
can be run without the necessity of any repairs. The
furnace cover lasts some twenty charges.

The labour cost is small, since three men can operate
a 12-ton unit. In this size the power consumption is some
800 kw. hours per metric ton of steel when starting up
from cold materials.

222 INDUSTRIAL ELECTROMETALLURGY

The carbons, of which there are four pendent ones, are
35 cms. in diameter, connected in parallel, and the furnace
operates at 70 to 75 volts with a current of 4000 amperes

12 ton Girgd furnace J.
Fig. 23. — Conducting hearth furnace for steel production.

per carbon. The average carbon consumption is about

6 kgms. per ton of steel.

I^ess load fluctuation is obtained in this type of arc
furnace, but it would appear
that the furnace lining has to
stand severer treatment than
that which obtains in the Heroult
or Stassano series arc types.

The Heroult and Keller fur-
naces are of the series arc type,
in which the current passes
from one electrode to the other
through the metal, striking arcs
between metal and electrodes in
its passage. Frequently three
pendent electrodes are used for
three phase-current, whilst Keller

has used four electrodes for simple alternating current.

The use of six electrodes in one hearth for three-phase

current has been suggested.

The furnace follows the normal construction, consisting

ip ■»„, H^nulr EU-«.

IRON AND THE FERRO-ALLOYS 223

of a steel shell with a magnesite or dolomite lining. The
roof liner is frequently made of silica brick. Since no
hearth electrode is employed, the liner may be further pro-
tected by a magnesite slag mixture bound with pitch.

The voltage lies between 90 and 100 volts, and the power
consumption per metric ton of steel produced with a cold
charge is from 700 to 800 kw. hours, and with a hot one
from 200 to 300 kw. hours. The electrodes are usually very
large, to reflect the arc down on to the surface of the slag
and thus protect the roof liner ; up to 60 cms. diameter
electrodes have actually been used. The electrode loss
is naturally heavier than in the Girod type, being about
12 kgms. per ton of steel produced.

These furnaces have been put to a variety of uses.
At Chicago, 15 Bessemer converter steel is blown until the
carbon and silicon are practically all removed. The metal
is then poured into the electric furnace, and .lime and iron
ore are added to remove the phosphorus. At the end of
half an hour the furnace is tilted, the slag removed, and a
fresh flux of lime, fluorspar and coke is added to remove
oxygen and sulphur. When the removal is complete, the
suitable amounts of carbon, ferrosilicon and f erromanganese
are added, the furnace is again tilted, and the metal run
from the ladle into the moulds.

At Syracuse, phosphorus and carbon are removed in
an open hearth furnace and the molten metal subsequently
transferred to a H6roult furnace for desulphurization.

At I<a Praz, three slags are formed and removed before
the final addition of the requisite amounts of ferro-alloy and
carbon are made to the steel.

Each slag removal necessitates the supply of an
additional 50 to 60 kw. hours per ton of steel produced.

Stassano's Furnace. — Captain Stassano in Italy was
one of the first investigators into the possibility of smelting
iron ores in the electric furnace. After a series of experi-
mental runs at Cerchi, large electric smelting plants were
installed at Darfo and Turin. He endeavoured to produce
steel in one operation directly from the ore. It is evident

224 INDUSTRIAL ELECTROMETALLURGY

that for the further refinement and decarburization of the
pig iron usually produced, provision must be made for the
supply of only the requisite amount of carbon and no more.
In addition the jnolten pig must be retained in the furnace
in such a maimer that the heterogeneous metal slag reactions,
by which the actual process of purification is accomplished,
have time to complete themselves.

Stassano accomplished the first by careful analysis of
the high-grade ore employed and briquetting it with the
requisite amount of carbon and flux, using pitch or water
glass as a binding material. The furnace itself consists of
a magnesia-lined cylinder with a domed roof capable of
slow revolution around a nearly vertical axis. Horizontal
electrodes, three or four in number, are employed, being
diametrically, introduced at the centre of the furnace cavity
and slightly inclined to the horizontal.

The briquetted charge is introduced at the top of the
furnace, and two tapping holes are provided for the with-
drawal of the metal and slag. At the commencement of the
operation a short arc is employed, but as the temperature
within the furnace rises the arc gap becomes more conducting
owing to the volatilization of impurities, and the electrodes
are withdrawn until an arc of some 40 to 50 cms. long is
obtained. Since the arc does not make any contact with
the ore or metal, heating is accomplished by radiation alone.

With a magnetite ore containing 48 to 50 per cent, of
iron 1 metric ton of metal could be produced with an energy
expenditure of 4800 to 5900 kw. hours, and an electrode
consumption of 10 to 15 kgms.

Various analyses of the resulting metal have been given 16
both by Stassano and by other investigators. The following
may be taken as the extreme limits of the carbon content : —

Per cent, composition.

(1)

(2)

Carbon

. . 0*80

0*090

Manganese

.. 0*30

0*092

Silicon

. . 0*22

Trace

Phosphorus . .

.. 0*015

0*009

Sulphur

.. 0-045

0*009

IRON AND THE FERRO-ALLOYS

225

The process has not extended beyond the confines
of Italy.

Induction Furnaces. — The Kjellin Furnace. This
furnace consists essentially of a step-down transformer in
which the secondary winding is replaced by an annular
trough of refractory material containing the molten steel,
A, A. This is excited by the primary B, B, and the lines
of force are retained as far as possible in the system by
the thin sheet-iron laminated core C, C.

The first furnace of this type was installed at Gyringe
in Sweden. With a primary alternating current of 90
amp&res at 3000 volts, the estimated induced current

WV,

Fig. 25. — Kjellin induction furnace.

was 30,000 amperes at 7 volts. The power consumption
was found to be 800 kw. hours per ton when charged
with cold metal, and 650 kw. hours per ton when charged
hot. Lindblad uses the following formula for determining
the power factor : —

V p* ) Is VW'^WV

where />=the power factor.

ft=the frequency.

^--(the ratio of area to length of the steel in the
/ J channel.

s=the sp. resistance of the steel.

C=a constant.
W 5 and W*=the magnetic resistances of the two circuits.

1,. 15

226 INDUSTRIAL ELECTROMETALLURGY

The power factor is consequently greatest when the
right-hand term is small. It would therefore appear
necessary to have a very low frequency current employing
a high secondary resistance in the form of a long thin trough.

The furnace referred to operated on a current of fre-
quency 13*5 cycles per second, having a power factor as
low as 0*635. A further disadvantage is to be found in
the fact that the secondary cannot be completely emptied
of metal, if it be desired to keep the furnace warm prior
to the insertion of a fresh charge. If a hot charge be placed
in the furnace its capacity is considerably augmented.

In the Colby and Gronwall furnaces, these difficulties
are partly overcome. Colby utilizes a water-cooled coiled
pipe as primary circuit, permitting of it being placed in
closer proximity to the secondary. A power factor of 0*90
to o*93 is claimed, and the calculated power comsumption
per ton of steel is 590 kw. hours for a cold charge and
490 kw. hours for a hot one. In Gronwall's furnace, a long
serpentine trough is used in order to ensure a high resistance
in the secondary circuit ; a high power factor is claimed.

In spite of the disadvantages of the simple induction
furnaces such as the low power factor with currents of
normal frequency, together with the difficulty of pro-
tecting a long trough of molten metal from excessive heat
radiation, several of the Kjellin type have been employed,
usually for the preparation of special steels and ferro-alloys
in which simple fusion operations are required, where local
overheating is to be avoided, and no slag formation is desired.
The slag is not usually sufficiently heated, and its removal
from the annular trough is a matter of considerable difficulty.

Frick furnaces, which are simple modifications of the
Kjellin, are in operation at Krupp's works at Essen for the
production of ferromanganese and melting scrap.

The following figures have been published relating to
these furnaces : —

Kw. hrs. per ton.

Ferromanganese production . . . . 600

Melting scrap 587

Steel refining 90

IRON AND THE FERRO-ALLOYS

227

Composite Furnaces. — The most successful composite
furnace employed in the preparation of steel is the Rochling
Rodenhauser resistance induction furnace.

In these furnaces the laminated soft iron cores A, A,
with the primary windings B, B, are surrounded by the
secondary molten-metal troughs D which are protected by
the magnesia-lined fireclay walls CE. The troughs meet in a
common space between the two cores, and a large reservoir
of molten metal is thus provided.

The heating of the metal in the trough is provided by
means of the induced current, but an extra supply of energy

IZZZZZZZ

(/,////

7JZ-L

UV/////V

////// /t

/ /Y//<

Fig. 26. — Rochling Rodenhauser resistance induction furnace.

has to be supplied to maintain the reservoir at the desired
temperature. This is accomplished by means of a few
turns of heavy cable wound round the pole pieces and
connected to iron plates F, F, embedded in the magnesia
liner E, E, at the opposite ends of the trough. The magnesia
becomes sufficiently conducting at high temperatures to
permit of the passage of the current induced in the cable
through the molten steel. About 65 per cent, of the in-
duced current goes through the annular troughs, and the
remaining 35 per cent, through the central reservoir.

Furnaces of this design have been built up to 8 tons
capacity, and have been found suitable for both pre-
paring and refining steel. The difficulties associated with
the electromagnetic rotation of the molten metal when

228 INDUSTRIAL ELECTROMETALLURGY

three-phase current is employed have already been
referred to.

To overcome the chief objection raised against this
furnace, viz. the low temperature of the slag, small arc
electrodes have been suggested as supplementary slag
heaters, as in the Paragon and Nathusius furnaces, the
requisite power being naturally obtained by an additional
secondary winding on the cores.

Although more expensive to erect, the induction furnaces
offer considerable advantages over the arc type, inasmuch as
no expense is entailed for carbon renewal, perfectly gas-free
metal can be obtained, and no impurities from carbon ash
are dissolved by the metal. The wear on the lining due
to the electromagnetically produced spin in the metal is
somewhat heavy.

Miscellaneous Furnaces. — Resistance furnaces such as
those of Gin n have not proved suitable for steel refining,
owing to the very high currents employed necessitated by
the low resistance of the molten metal.

Two interesting types of furnaces which have not yet
been applied on a technical scale may be mentioned, since
the application of the principles employed are novel for the
purpose in view, and laboratory experiments have yielded
highly satisfactory results.

The Hering "Pinch" Effect Furnace. 1 * — Hering, when
investigating the operation of the Kjellin furnace, noticed
that when high current densities were employed the surface
of the metal became occasionally depressed ; ultimately the
ring was divided and the current ceased. He pointed out
that the depression was caused by the pressure directed
towards the axis caused by the mutual attraction of the
coaxial cylinders of metal carrying the induced currents.
Northup has shown that this pressure exerted perpendicu-
larly to the axis of the cylinder is proportional to the square
of the current and inversely to the square of the radius.
Consequently any slight difference in diameter of the fluid
metallic conductor will produce a great alteration in the
axial pressure. Under these conditions the fluid will be

IRON AND THE FERRO-ALLOYS 229

forced from the constricted area, thus increasing the " pinch "
effect, and a ruption will ultimately result Hering has
applied the "pinch" effect to a small-scale furnace with
success.

The molten metal contained in a reservoir is connected
to water-cooled electrodes through two narrow channels
containing some of the molten metal, Tyhich in turn are
connected to the secondary winding of a transformer.
Very active circulation of the metal is caused by the con-
tinuous " pinching " of the metal in the tubes. Heating
accomplished by the passage of the current from the
secondary, according to Hering, is slightly augmented by
the frictional heating in the tubes. The furnace has been
employed successfully as a crucible furnace for steel melting,
and preliminary experiments have been made on the direct
production of pig iron.

It is evident that the furnace thus designed should
possess considerable advantages over the ordinary in-
duction furnace. The rapid circulation of the metal en-
sures a uniform temperature distribution, and should con-
siderably accelerate slag metal reactions owing to the
continuous renewal of the surface of contact. High slag
temperatures are more easily obtained owing to the fact
that very thick furnace walls can be used.

Although the wear on the " pinching " channels is
liable to prove excessive, and possible emulsification of the
slag in the metal may occur in the channels themselves,
large-scale experiments on furnaces of this design would
probably give results better than those of the ordinary
induction furnace, and certainly better than those given
by the resistance furnaces.

Northrup's Tesla Induction Furnace. — Northrup ld has
pointed out that the limitations of the ordinary induc-
tion furnace are determined by the " pinch " effect in the
fluid secondary winding, and the excessive magnetic leakage
in the usually accepted annular form of construction.
He has accordingly designed a crucible furnace thermally
and electrically insulated on the outside ; this is wound

230 INDUSTRIAL ELECTROMETALLURGY

with about fifty turns of wire, which serve as the primary of
the induction coil. The ends of this primary are connected
to the electrodes of a Tesla coil fitted with condensers and
capable of providing very high voltage oscillating discharges.
The metal in the crucible serves as the secondary of the
coil. A 20-kw. furnace has been constructed and found to
operate successfully with a condenser terminal voltage of
5400 to 7200 volts, and the natural period of oscillation of
the discharge. The thermal efficiency is stated to be 60 per
cent., a high figure when the small size of the furnace is
considered.

C. The Ferro-au,oys.

The production of ferro-alloys in the electric furnace was
one of the earliest applications of electrothermal methods
to the preparation of iron and steel. Amongst the most
important alloys manufactured may be mentioned ferro-
silicon, ferro-tungsten, manganese, chrome, molybdenum
and smaller quantities of ferro-uranium and titanium.

Ferro - silicon. — Arc furnaces are generally employed
for the production of ferro-alloys, either with a basal elec-
trode such as the Hfroult, in which combined arc and
resistance heating are employed, or the series arc type as in
the Keller (p. 212).

In the preparation of ferro-silicon originally iron ore
was used, but scrap iron is now employed. It is made in
several grades, containing 25 per cent., 50 per cent., 75 per
cent., and over 90 per cent, silicon. The preparation of
the purer silicon grade has already been described.

Ferro-silicon should be prepared from scrap iron of
low phosphorus content, since the presence of calcium
phosphide has been shown to be the source of explosions
and cases of poisoning, formerly of frequent occurrence in
the manufacture and handling of the substance. Ferro-
silicon containing over 70 per cent, silicon is more stable than
the lower grades. The raw materials used are crushed quartz
and carbon in the form of anthracite or coke. Sand has also
been experimented with, but is liable to choke the furnace.

IRON AND THE FERRO-ALLOYS 231

The furnace charge crushed to a small size should contain
sufficient carbon to reduce the quartz, according to the

equation- S i0 2 + 3 C=2CO+Si

and iron is added in amount depending on the grade of
ferro-silicon required.

The voltage employed with a single-arc furnace varies
from 70 to 75 volts, and the power consumption for a 75
per cent, grade ferro-silicon is roughly 5000 kw. hours per
metric ton, for a 30 per cent, ferro-silicon only 3500 kw.
hours are necessary.

About 80 per cent, of the charge is converted into
utilizable ferro-silicon; the remainder is used for slagging
off the impurities in the quartz and coke. Ferro-silicon
absorbs very little carbon during the process of formation,
and furnace liners of carbon are frequently employed.

Attempts, partly successful, have been made to utilize
blast furnace slags, 20 and ordinary sandstone 21 rock as
source of the silica.

An application of the electric furnace has recently been
made to the preparation of special ferro-silicon and silicized
iron having resistant properties, probably associated with the
formation of superficial layers of iron silicides, FeSi* and FeSi2-

Owing to the stimulus given by the war to the production
of strong acids, a great number of these non-corroding
castings have been introduced under a variety of names, such
as Tantiron, Narki, Illenit, Neutraleisen and Metaldtir.
The earlier forms were exceedingly brittle, and could not
be machined, but recently large castings capable of being
machined have been introduced, and the presence of flaws
practically eliminated.

Ferro - tungsten. — Ferro - tungsten is generally pre-
pared on the intermittent system. The charge is fused in
a simple furnace lined with clay having a pendent and one
basal electrode. When the reduction is completed, the
charge is allowed to solidify, and is then broken out.
Attempts have also been made to use tilting furnaces to
avoid the time wasted in cooling the melt.

232 INDUSTRIAL ELECTROMETALLURGY

As source of tungsten, various ores and ore concentrates
are used, the most common being scheelite, CaW0 4 . Re-
duction is usually accomplished by means of coke, and the
iron is supplied by the addition of hematite. Sulphide
of iron has also been used : 22

CaW0 4 +FeS+4C=(Fe,W) +CaS+4CO

Gin has proposed the use of ferro-silicon as a reducing
agent instead of carbon, but the process does not appear
economical —

3CaW0 4 +4Fe2Si =3CaSi0 3 +FeSi0 3 + (Fe, W)

Ferberite, Fe 2 W0 4 , and wolframite, FeMnW0 4 , are other
important tungsten ores, and can be directly reduced with
carbon in the electric furnace, most of the manganese being
volatilized at the temperature of reduction, 2800 C. The
loss of tungsten in the operation of reduction is usually
small, but decarburization of the alloy is usually essential
owing to the formation o£ tungsten carbide, W 2 C.

Decarburization for low-grade tungsten alloys can most
easily be accomplished by the addition of a strictly limited
amount of oxide of iron. Excess of iron oxide is to be
avoided owing to the formation of feirous tungstate. For
higher grades, calcium carbide and ferro-silicon with a flux
of calcium fluoride are employed, any silica present in the
concentrates being fluxed by the addition of lime.

According to Keeney 28 the power consumption for
reduction and decarburization can be reduced to under
7500 kw. hours per metric ton. Hutton 24 gives the
following analyses of two typical industrial alloys : —

Percentage composition.

(1)

(2)

Tungsten . .

.. 85-15

71-80

Iron

. . 14*12

24*35

Carbon

. . 0-45

2-58

Silicon

. . 0-13

0-36

Manganese . .

. . 0-085

0'75

Sulphur

. . 0'02I

0*02

Phosphorus

. . o # oi8

0-008

IRON AND THE FERRO-ALLOYS 233

Metallic tungsten and high-grade tungsten alloys are
used for the production of crucible tool steels, whilst the
lower grades with the higher carbon content are employed
for open -hearth steels containing low percentage of the
metal.

Ferro-manganese. — Ferro-manganese is usually pre-
pared in the blast furnace, 25 but different grades of the alloy
are prepared by the fusion of scrap metal and manganese in
the electric furnace. As has already been indicated, induc-
tion furnaces (p. 225) appear most suitable for this work,
although Heroult and Girod furnaces have been employed
for the purpose.

To prevent absorption of manganese by the calcined
dolomite liner, the walls are frequently protected with a tar
or a mixture of retort coke and coal tar.

Ferro-chrome. — Ferro-chrome has found an increasing
field for use in special steels for naval and military purposes,
and also in the production of the so-called " rustless "
steels, which, although malleable and capable of being
welded, are resistant to sea- water and acids.

A continuous operating furnace can be employed, being
tapped at the base. The furnace walls are usually lined
with dolomite or magnesite, but frequently a chromite
liner is employed ; with careful operation, the life of a liner
may exceed three years. The chief source of chromium
is the mineral chromite, FeO.Cr 2 3 , and reduction is
usually accomplished by means of carbon, although silicon,
aluminium and calcium carbide have been suggested.
The latter processes have not proved economically suc-
cessful.

The charge of finely powdered chromite and coarse
anthracite or coke in the requisite quantities to ensure
reduction is fed into the furnace at regular intervals.

Reduction commences at about 1185 C. 26 By inter-
mittent tapping a ferro-chrome containing only from 2 to 5
per cent, of carbon can be run off, although the carbon content
may run considerably higher. For the purpose of preparing a
low carbon f errochrome, a decarburizing process is necessary,

234 INDUSTRIAL ELECTROMETALLURGY

since the direct production of a low carbon alloy is,
according to Keeney, 27 always attended by an excessive loss
of chromium in the slag. Refining is usually accomplished
by fusion of the alloy with a suitable slag containing chromite
or oxide of iron, lime and fluorspar. Decarburization is
accomplished according to the following equation : —

2Fe 3 C+6Cr a C3+5FeO.Cr2O3=iiFeCr 2 +20CO

The carbon of the resulting alloy is usually below 0*5 per cent,
and may be lower.

Hutton gives the following analysis of commercial
ferro-chromium : —

Percentage composition.

Iron

.. 2705

Carbon

.. 425

Silicon

. . o-6o

Manganese

. . 046

Aluminium

. . 0*22

Magnesium

. . 0-3I

Sulphur

. . 0*02

Phosphorus

. . 0*02

The power consumption ranges between 6000 and
7200 kw. hours per metric ton, with an electrode loss of
25 kgms. of carbon.

Chromium-nickel alloys for the production of high speed
tool metal have recently been introduced and are being
prepared in increasing quantities.

Ferro-molybdenum. — Ferro-molybdenum is prepared
from molybdenite ore or concentrates averaging 90 per cent.
MoS 2 . The correct proportions of iron turnings, anthracite,
coal or coke and the raw or roasted ore, together with lime,
are heated in an intermittent electric furnace, usually of
the basal electrode type. Reduction takes place according
to the equation —

2M0S2 +2CaO +3C +Fe =FeMo 2 +2CaS +2CO +CS 2

The resulting alloy usually contains from 3 to 4 per cent,
of carbon. A typical analysis is as follows : —

IRON AND THE FERRO-ALLOYS

235

Percentage composition;

Molybdenum

. . 80*20

Iron

12*65

Sulphur

0*025

Phosphorus

0*028

Carbon

327

Decarburization and desulphurization can be accom-
plished by means of a slag containing lime and oxide of
iron slag. Reduction by means of carbides, aluminium and
silicon, including ferro-silicon, have all proved too expensive
for commercial practice. The loss by volatilization of
molybdenum oxide is frequently very high and may amount
to as much as 30 per cent.

Ferro - vanadium. — The electric furnace method for
preparing ferro-vanadium has only recently supplanted the
more usual thermite process. A great variety of processes
have been suggested for the production of the alloy, amongst
which may be mentioned —

1. Fusion of a mixture of 10 parts vanadium pentoxide,
1 part of silica and 3 parts of carbon, with the requisite
amount of iron.

2. Briquetting vanadium trioxide and ferro-silicon by
means of tar.

3. Electrolysis with an iron cathode from a double
fluoride electrolyte.

4. From ferro-vanadium silicide, SiFeV and vanadium
fluoride.

5. From the oxides, reduction being brought about
by means of carbon. A current of 900 amp&res at 50 volts
in a small arc furnace will provide an alloy containing from
4 to 6 per cent, carbon. By reheating with a limited amount
of oxide the carbon content can be reduced to under 1 per cent.

The largest source of supply is the sulphide ore, patronite,
and experiments by Keeney have shown that the preparation
of the ferro-alloy can be accomplished in a manner similar
to that employed for ferro-molybdenum.

Ferro-titanium, Uranium and Boron.— Ferro-titan-
ium, used as a deoxidizer for cast iron, is made by smelting

236 INDUSTRIAL ELECTROMETALLURGY

titaniferous iron ore with carbon or aluminium, whilst the
uranium alloy has been prepared in small quantities from
sodium uranate, Na 2 Ur 2 7 , or uranium oxide, U 3 8 , by
smelting with iron sulphide and lime or with oxide of iron
and calcium carbide or ferrosilicon. The boron alloy is
prepared by reduction of a mixture of scrap iron and borax
or boric acid with carbon.

REFERENCES TO SECTION VIII.

1 " The Rusting of Iron and Steel," by E. K. Rideal.
3 Zeit. Elehtrochem., 16, 20; 1910.

3 D.R., patent 859 of 1900.

4 Trans. Amer. Electrochem. Soc., 19, 181 ; 191 1.

* Trans. Amer. Electrochem. Soc., 25, p. 489 ; 1914.
Trans. Amer. Electrochem. Soc, 25, 1914. P#*"i^

7 " Quantitative Analyse durch Elektrolyse," 5th Auf, p. 172 ; 1908.

8 J. Richards, Electrochem. Ind., 7, 16; 1907. Wright, "Electric

Furnaces." Allmand, " Applied Electrochemistry."

• J. Crawford, Met. and Chem. Eng., 11, 1913. P- 3 8 3»

10 Askenasy, " Einfiihrung in die technische Elektrochemie." 191 o.

11 See " Congress of Applied Chemistry." 7 ; 1912.
13 Chem. Zeit., 86, p. 564 ; 1912.

13 Trans. Faraday Soc, Dec., 191 6.

14 Trans. Amer. Electrochem. Soc, 15, 127 ; 1909.
14 Stansfield, " The Electric Furnace."

13 Electrochem: and Met. Ind., vol. 6, 1908, p. 315 ; vol. 9, 1911, p. 642.

17 Trans. Amer. Electrochem. Soc, 15, p. 205 ; 1907.

18 Trans. Amer. Electrochem. Soc, 15, 255 ; 1909.

19 Trans. Farad. Soc, Nov. 7, 19x7.

30 G. Gin, Industrial Electro., April, 1901.

31 Met. and Chem. Eng., 8, p. 134 ; 1910.
33 Eng. and Min. Jour., 18, 173; 1912.

33 Trans. Amer. Electrochem. Soc, 24, p. 182 ; 1914.

34 " Electrochem. Industry," vol. 5, p. 10.

38 F. W. Harbord, " The Metallurgy of Steel."

36 J.C.S., 93, 1484 ; 1908.

37 Trans. Amer. Electrochem. Soc, 24, p. 177; 19".

BIBLIOGRAPHY.

" Elektrische Ofen in der Eisenindustrie," Rodenhauser.

' ' Electric Furnaces and their Industrial Applications, ' ' S. Wright. 1 904.

" The Electric Furnace," A. Stansfield. 1914-

"Applied Electrochemistry," A. J. Allmand. 1912.

"The Metallurgy of Steel," F. W. Harbord.

" Einfiihrung in die technische ElektTochemie," Askenasy. 1910.

"Stahlu. Eisen."

APPENDIX

SOME ELECTROLYTIC PROPERTIES OF THE ELEMENTS

Element.

Potassium

Sodium

Barium
Strontium

Calcium . . .

Aluminium

Magnesium
Molybdenum
Chromium
Manganese

Zinc

Indium
Cadmium ...
Iron

Thallium ...

Cobalt
Nickel
Tellurium...
lin

Lead

Hydrogen . . .

Copper

Bismuth ...
Antimony...
Mercury ...
Silver
Palladium
Platinum ...

Gold

Electrolytic
Potential E k .

Electro-
lytic over-
voltage
q to H a .

+3*20

+2*82

+2-82
+277

+ 2*56

+ ?

+ 1-49

+ ?
+ ? *
+ 1*075

+0770

+045
+0*42
+034
+0*322
(in Tl so-
lutions)
4-0*232
+0*228

?
+0*192

+0*148

—0*329

-0393
-0463
-0750
-0771

-0793
—0-863

?
p

?
?

Electro-
chemical
equiva-
lent wt.
of metal
in grras.
deposited
per am-
pere-hour.

?
?
?
?

0*70

?
048
008

0*03-0*21

o*43-o*53

0*35-0*64

0*03-0*23

?
0-42-0*78
0*05-0*15
0*23-0*46
0*07-0*09

0*02-0*06

X'46

o*86

2*56
1 63

075

o*34

o*45
o*6o

065

1*03

Val-
ency
in
solu-
tion.

1*22

1*43
2*09

I'lO

3*8o

I*IO

1*07

2-38

2*22

3*86

0*0376

/2*37
\ii8

2*58

1*49
7*46

4*02

1*99
i*8i

|* # 45
\7*35

2
2

Usual current densities employed for

deposition, refining, plating in amperes

per sq. decimetre.

2
6

2
2

Deposition.

200

{Castner 200
Griesheim
1000

110-250, con-
tact elec-
trode up to
10,000
/Hall 100 1
\Heroult 190]
10-15

elec

trolyte
Chloride |

electro- 73-4
1 lyte )

1 -3-1-5

10
10

3
3

I
2

0*5-2

0*005-0*01

Refining.

Plating.

0-8-1

1*4
3-20

i-o

2-3*5

o'8-i

o*3-4
03

1-2 O'H

rSilico-

fluoridel

0*9-2*2
iPerchlo-

ride 2-3 J

0-5-2*5

01-0*4

NAME INDEX

ACHESON, 155, 165, 169

Acker, 117
Allmand, 166, 173
Anderson, 85
Andrioli, 72
Arndt, 121
Arrhenius, 2, 3
Aschermann, 151
Ashcroft, 60, 115, 123

Bancroft, 53

Bassett, 69

Bayer, 126

Beardslee, 104

Beatson, 93

Becker, 112

Becket, 152

Benjerink, 191

Bennet, 65

Bergsoe, 95

Berkeland, 186

Berthelot, 196

Bessemer, 28

Betts, 85, 87, 91

Beutel, 74

Bicknell, 121

Bischof, 102

Blount, 72, 123, 132

Body, 37

Boissiere, 199

Borchers, 28, 94, 120, 122, 123, 137,

146, 150, 155. 179
Bottger, 74, 104
Brand, 68
Broadrill, 137
Brochet, 101
Brode, no
Brown, 94, 146
Brunner, Mond, 65
Bucher, 196, 199
Bullier, 177
Bunsen, 138
Burgess, 206
Burleigh, 86

Carmichael, 34

Caro, 200

Carrier, 117

Caspari, 8, 100

Castner, 109, 196

Cavendish, 186

Clancy, 72

Clarke, 199

Classen, 44, 207

Claude, 201

Claus, 92

Clergue, 150

Cohen, 73, 169

Colby, 227

Consiglio, 42

Cowles. 133, 144

Cowper Coles, 42, 43, 72, 206, 208

Crookes, 183

Cullis, 22

Daniel, 51
Darling, 112
Davy 10, 186
Dawes, 199
Dechert, 102
De Laval, 147
Dennis, 135
Desmur, 102
Deville, 119, 154
Diesel, 20
Doelter, 220
Dolch, 93
Dorsemagen, 146
Dumoulin, 42

Elkington, 82
Elmer, 97
Elmore, 42
Eisner, 74, 83
Englehardt, 95

Faraday, r, 10
Feild, 219
Field, 97

240

NAME INDEX

229,

Fischer, 49, 69, 88, 117
Fitzgerald, 146, 155, 168
Foerster, 66, 82, 92, 93. *35
Forsell, 155
Frank, 200
Frary, 121
Frick, 226
Fromm, 62, 66

Gaudin, 44

Geer, 135

Gelsthorpe, 94, 95

Gillet, 165

Gin, 130, 137, 142, 150

233
Girod, 220, 234
Goldschmidt, 93
Goodwin, 121
Greenawalt, 37
Greenwood, 153
Griesheim, in
Gronwall, 214, 227
Grosz, 154
Grotthus, 2
Gruszkrewicz, 197
Guichard, 152
Gunther, 66, 98
Guntz, 138

Haanel, 213

Haber, 175, 187, 188

Hall, 125

Hambuechen, 206

Hampe, 137

Harbord, 142

Harden, 150

Harmet, 214

Harper, 103

Hausser, 188

Haycroft, 72

Helfenstein, 179, 214

Hellriegel, 192

Helmholtz, 5

Hemingmay, 95

Henderson, 5

Hering, 159, 229, 230

Heroult, 125, 213, 222, 231, 234

Hewes, 177

Heyn. 149

Hildebrand, 136

Hittorf, 1, 45

Hoepfner, 35, 65, 99

Holfis, 96

Horry, 177

Howies, 186

Hnlin, 117

Hutton, 233, 235

Imbert, 146
Irvine, 158

Jacobsen, 73
Jellinek, 186
Johnson, 141, 145

Kalmus, 103

Keeney, 152, 233, 235, 236

Keith, 86, 94

Keller, 148, 213, 222, 231

Kendall, 198

Kern, 88, 98

Kjellin, 226

Klapproth, 76, 90

Koenig, 187

Kohlrausch, 1

Krupp, 227

Krutwig, 83

Kuster, 81

Ladd, 149

Lampen, 165

Landis, 158

Langbein, 73, 82, 102, 104, 206

Laszczynski, 33, 59

LebJanc, 9, no, 201

Leaner, 152

Lepieme, 135

Lepinske, 197

Leucks, 87

Li, 207

Lindblad, 214, 226

Linde, 201

Lodge, 2

Lorenz, 10, 124

Lowry, 187

Luckow, 94 '

Lyon, 214

Lyons, 137

Machalske, 158
Marchese, 28
Margueritte, 199
Mathers, 89
Matuscheck, 96
Maxted, 88
McDougall, 186
Meliner, 199
Memmo, 179
Mennicke, 96
Merke, 206
Minet, 130, 154
Moebius, 77, 79
Moissan, 150, 153, 172
Morrison, 149
Mounden, 142
Muthmann, 136
Mylius, 62, 66

NAME INDEX

241

Namias, 82
Nathusius, 222
Nauhardt, 95
Neil, 94
Neill, 34

Nernst, 15, 42, 47, 186
Neumann, 157, 217
Newton, 199
Nicola jew, 28
Nodin, 96
Northrup, 229, 230
Norton, 136
Noyes, 47, 48

Obstbrlb, 146
Oettel, 119
Ost, 76, 90
Overman, 89

Palmabk, 5

Parker, 158

Parkes, 59, 67, 86

Paschen, 5

Pattison, 86

Pauling, 186

Pederson, 137

Perkin, 76

Pfanhauser, 74, 82, 101, 102, 201

Planck, 5

Plato, 120

Possoz, 199

Pott, 102

Potter, 154, 169

Powell, 102

Preeble, 76

Priag. 63, 77. 8 7» 128

Quintains, 95

Ramsay, 188

Rayleigh, 186

Read, 193

Readmann, 158, 198

Regnault, 188

Richards, 67, 130, 139, 159, 166

Ricketts, 33

Rienders, 95

Rochling Rodenhauser, 228

Roseleur, 73, 97

Rossi, 184, 187

Rudolphi, 173

Ruff, 120

Russell, 69

Salgues, 142, 144
Sand, 45, 81, 92
Saunders, 165
Savell, 103

I,.

Scholl, 113

Schonherr, 186

Schucht, 135

Schwabe, 92

Senn, 88

Serpek, 192

Seward, 114

Sharpe, 65

Siemens Halske, 29, 59, 71, 90

Simon, 138

Smit, 5

Smith, 32, 76, 95

Snowden, 81, 89

Snyder, 125

Sourdeval, 199

Stalhane, 214

Stansneld, 142, 154* 214

Stassano, 222, 224, 225

Steele, 2

Steiner, 94

Stockem, 120

Strutt, 187

Suchy, 124

Swan, 198

Swinburne, 123

Swindel, 199

Tainton, 63
Taylor, 168
Tesla, 231
Thiel, 135
Thiele, 88
Thompson, 65, 175
Thomson, 5
Thum, 79
Tommasi, 89
Tone, 166, 168, 170
Townsend, 155
Tronson, 121
Trumm, 98
Tucker, 120, 165, 193
Tuttle, 74

Valentine, 85
Van Arsdale, 34
Van der Waal, 3
Van Laar, 5
Van *t Hoff, 2, 3
Vautin, 117
Vogel, 124
Vogt, 220
Von Hevesy, no
Von Kugelgen, 114
Von Ruolz, 76

Waldbn, 3
Wallace, 76
Wannschaft, 99
Watt, 44

242

Watts, 65, 103, 104, 207
Weidlein, 34
Weintraub, 170
Westman, 159
Whetham, 2
Whitney, 47, 48
Wilson, 177

NAME INDEX

Winogradsky, 191
Withrow, 76
W6hler, 120
Wohlwill, 79
Wood, 136
Woolrich, 69
Wright, 132

SUBJECT INDEX

Abrasives, 164, 169, 173
Absorption, 8
Acetate electrolytes, 102
Active hydrogen, 65
Addition agents, 52
Alcohol reducers, 53
Algae, 192
Allotropy, 9, 187
Alloys, 54
Aluminium, 8

alloys, 133

carbide, 193

cathodes, 62

deposits, 26

nitride, 193

preparation, 125

reduction, 218, 234, 237
Alumite, 26
Alunduxn, 157
Ammonia, 157, 184, 186
Analysis, 52
Anodes, 30, 60, 131
Anthracite, 155, 157
Antimony, 39, 84, 89, 92, 151
Apatite, 158
Aquadag, 157
Arc furnaces, 153, 186, 198, 221,

222
Argentium, 133
Argon, 190
Aromatic acids, 208
Arsenic, 39, 40, 84, 88, 89, 159, 176
Arsine, 176

Asymmetric alternating current, 76
Atomistic theory of electricity, x
Auric salts, 71
Aureus salts, 72
Auxiliary electrodes, n

BACILLUS radicMa, 191, 192
Bacteria, 191
Barium, 122

cyanamide, 201

cyanide, 198, 199
Basic salts, 63
Bauxite, 22, 125, 193

Benzoic acid, 102

naphthol, 65

Biochemical nitrogen, 191

Bipolar electrodes, 31

Bismuth. 39, 79, 81, 84, 88, 89, 91

Blast furnaces, 177

Bleaching powder, 60

Blende, 59

Block furnaces, 208

Blowing off waste gas, 190

Blue powder, 59, 140

Boiler efficiency, 18

Bombs, 190

Boric acid, 102, 105, 237

Borides, 181

Boron, 137, 195

nitride, 195
Brass, 54
Brighteners, 83
Briquetting, 59, 225
Broken Hill ore, 59
Bromide electrolytes, 76
Bronze. 54, 133
Burnishing, 50
By-products, 13

Cadmium, 8, 68, 84, 135
Calamine, 65
Calcium, 120

carbide, 152, 234

cyanamide, 72, 200, 204

fluoride, 120, 220, 233

phosphide, 2x9, 231
Caliche, 184
Carbon, 39

disulphide, 83, 146, 160
Carborundum, 147, 164, 170
Carnallite, 1 17, 121
Cast iron, 213

Catalysis, 155, 190, 193. 198, 199
Cataphoresis, 11
Cathode depolarisation, 54

material, 61

potential, 99

rotation, 45
Cerium, 136

244

SUBJECT INDEX

Channel formation, 146

Charcoal, 209, 214

Chemical potential, 4

Chemical side reactions, 10

Chilled arcs, 187

Chloride removal, 33

Chlorine evolution, 66, 124

Chrome nickel alloys, 235

Chromite, 23, 234

Chromium, 8, 23, 151, 198, 218, 235

Citric acid, 102, 107

Cloud formation, 10, 127

Coal, 17

tar, 19, 234
Cobalt, 4, 23, 84, 103, 189
Coconut matting, 32
Collodion, 83

Colloids, 12, 52, 63, 65, 81, 87
Combustion, 188
Comminuted charges, 175
Complex electrolytes, 44, 55
Contact electrc : js, in
Copper, 25, 28, 39, 41, 79, 81, 147

hydride, 54
Corrosion, 61
Crucible steel, 239
Cryolite, 125, 132
Cuppelation, 77
Cupramines, 51
Cupriferous pyrites, 25
Cyanamide, 192, 200, 204
Cyanides, 183, 192

electrolytes, 42, 82

gold process, 70
Cyanogen, 196

Decarburization, 233, 234, 235,

236
Deoxidation, 218
Dephosphorization, 165, 218
Depolarizers, 7

Destructive distillation, 176, 185
Desulphurization, 219, 236
Dextrin, 65
Diamond, 154
Diaphragms, 5, n, 32, 79
Diffusion currents, 45, no

films, 47
Dissociation theory, 2
Dolomite liners, 213, 224, 234
Dropping electrode, 5
Duralium, 133
Dyes, 183
Dynamic equilibrium, 45

Education, 21
Eikonogen, 65
Electrode arrangements, 39
preparation, 157

Electrolytic agitation, 45

potentials, 3
Electromagnetic fields, 116, 222, 229
Electronic theory, 1, 187
Electroplating, 41, 52, 81, 89
Electrostatic charges, 140
Electrotype, 41
Emulsification, 222
Engines, 20

Equilibrium constant, 1S8, zoo
Evaporation, 10
Explosives, 183

Felspars, 21
Ferberite, 233
Fermentation, 18, 185
Ferric chloride leach, 37
Ferro alloys, 153, 208

boron, 237

chrome, 151, 234

manganese, 150, 226, 227, 234

molybdenum, 235

silicon, 35, 146, 153, 220, 231

titanium, 236

tungsten, 232

uranium, 236

vanadium, 236
Ferrocyanides, 73, 74
Ferrous sulphate, 91

depolarizer, 60

tungstate, 233
Fertilizers, 183
Fire sand, 168
Flashing, 42
Float slimes, 40
Flotation of ores, 59
Fluoborate electrolytes, 102
Fluorspar, 224, 235
Fluosiucate electrolytes, 85, 92, 102
Fluxes, 152, 208
Fog formation, 120, 131
Food supply, 183
Fractional crystallization, 86
Furnaces,

arc, 153, 186, 222

induction, 226

radiation, 147

resistance, 228
Fused electrolytes, 10, 109

Galena, 25, 59
Gallium, 135
Gallo tannic acid, 157
Galvanizer's dross, 57
Galvanizing, 59, 67
Garnet, 25
Gas films, 8

power, 18, 20

producers, 189

SUBJECT INDEX

245

Gelatine, 75, 80, 84, 87, 103

Glucosides, 103

Glue, 83, 87, 103

Glycerine, 113

Gold. 39, 69, 84, 71, 74, 75, 79, 84,

159
Graphite, 151, 154, 165

anodes, 97
Gredag, 157
Guilds, 184
Gum, 65
Guttapercha, 83

Heat insulators, 169, 170, 209
Hematite, 22, 209, 214, 233
Heterogeneous reactions, 217, 219,

222
Hexaminetetramine, 206
Homogeneous reactions, 217, 219,

222
Hydration of ions, 53
Hydrides, 8, 63
Hydrocarbons, 173, 198
Hydrocyanic acid, 196
Hydroelectric power, 15
Hydrofluoric acid, 87
Hydrogen electrode, 5

evolution, 65, 104

purification, 189
Hydroxylamine, 53

Illbnit, 232

Impurities in copper, 39

Inclusion of electrolyte, 40

Indium, 135

Induction furnaces, 221, 230

Ingot furnaces, 177

Iodides, 71

Ionic velocities, 1

Iridium, 75

Iron anodes, 60, 71, 207

deposits, 22

electrolytes, 206

electrothermal, 208

impurities, 39, 59, 66, 89

Labile hydrates, 53
Lactates, 98
Lanthanum, 136
Laterite, 26
Lead anodes, 30, 60

cathodes, 71, 96

chloride, 85

deposits, 25

impurities, 39, 59, 81

plating, 89

refining, 83, 84, 122
Lime, 229, 233
Limestone, 23, 180, 210

I,.

Limonite, 209
Load factor, 13
Lubricants, 157

Macrocrystalline deposits, 69, 81
Magnalium, 133
Magnesia, 115, 222, 225, 228
Magnesite, 215, 224, 234
Magnesium, 117, 145

chloride, 118

nitride, 195
Magnetite, 22, 30, 60, 209, 214, 225
Manganese, 23, 66, 138, 150, 198,

209
Manganese oxide anodes, 30
Mechanical burnishers, 42

scrapers, 78, 87
Mercury, 70, 72
Metaldur, 232
Metal fog, 119
Methane, 194
Mica diaphragms, 35
Molasses, 53
Molybdenite, 23, 235
Molybdenum, 8, 23, 152, 198
Monazite, 21
Mottramite, 24

Narki, 232
Neodymium, 136
Neutraleisen, 232

Nickel, 24, 39, 4°. 54. 84, 98, 100,
102. 149

plating, 100
Nitre, 184
Nitric acid, 112

oxide, 186
Nitrides, 192

Nitrogen fixation, 186, 201
NitroUm, 202
Noble metals, 8

Oildag, 157
Ores, iron, 207
Organic solvents, 120
Oscillation discharge, 231

of ions, 9
Osmium, 75, 189
Overpotential, 6, 7, 61, 206
Oxalates, 209
Oxide films, 9
Oxides of nitrogen, 187
Oxidizing agents, 54
Oxycarbides of silicon, 165
Oxychlorides, 121
Oxygen, 39

Palladium, 75
Parting of gold, 77, 81

16*

246

SUBJECT INDEX

Passivity, 6, 8, 43
Patronite, 236
Peasant proprietors, 183
Peat, 185

Peeling of metal, 104
Peptone, 89, 193
Perchlorate electrolytes, 88
Phenol, 88

Phosphate electrotytes, 73, 97
Phosphine, 176
Phosphorus, 158, 191, 209
Pig iron, 208
Pigments, 109
Pinch effect, 229
Platinum, 8, 75
Polishing, 169
Porous deposits, 42
Potassium, 117
cyanide, 197
Power, 12, 15, 17, 18, 209, 217, 226
Praseodymium, 131
Printer's ink, 169
Protective colloids, 12, 52, 53
Prussian blue, 71
Pyrogallol, 53, 65, 88
Pyrometers, 156
Pyrrhotite, 214

Quartz, 25, 87, 231, 232

Radiation furnaces, 147
Reaction velocity, 9
Reducing agents, 53, 219
Refining, 37, 66

Resistance furnaces, 145, 175, 221
Resources of ores, 21
Revolving carbide furnace, 178
Rhodamite, 25
Roasting processes, 59
Rotation of cathodes, 45, 97
Rubite, 24
Rustless steel, 234

Samarium, 136
Self-induction, 175
Sewage sludge, 185
Shaft furnaces, 214
Sickening of mercury, 70
Silfrax, 168
Silica bricks, 219, 224

soluble, 59
Silicon, 153, 195, 220, 234

carbide, 165

monoxide, 165

nitride, 195
Silidizing, 168
Silit, 169

Siloxicon, 165, 176
Silundum, 168

Silver, 39, 79, Si, 84, 159

Slags, 143. 147, 150, 208, 217, 220

Slimes, 40, 77, 86

Smoke loss, 18

Smothered arc furnace, 175

Sodamide, 189

Sodium, 109, 113, 114, 191, 196

Sodium,

cyanide, 199

peroxide, 70

sulphide, 113
Softening Doint of slags, 220
Solid solutions, 40
Solution pressure, 3, 10
Soot, 197

Spark discharge, 198
Spelter, 140
Spiegeleisen, 150
Spongy deposits, 42, 65, 81
Stannates, 94
Stassfurt deposits, 117
Steel, 216

Striking baths, 82, xoi
Strontium, 122
Sublimation pressure, n, 166
Sugar, 53
Sulphide of carbon, 160

gold, 76
Sulphocyanides, 73
Sulphur, 39, 82
Sulphur dioxide, 35, 218
Sulphuric acid, 32
Surface tension, 8
Symbiosis, 192
Synthetic ammonia, 188, 196

Tannin, 96, 98, 103
Tantiron, 232
Tapping furnaces, 179
Tartrate electrolytes, 92, 98, 207
Tellurium, 89
Thallum, 135
Thermite, 153
Thioantimonates, 91
Thiostannates, 92
Tidal energy, 14
Tilting furnaces, 232

Tin, 26, 92, 96, 150

Tin scrap, 93

Titanium, 24, 137, 153. !95» 198
236
nitride, 195

Transmission of power, 15

Transport numbers, 1

Tungsten, 8, 153, 187, 218, 234
carbide, 233

Turbo-generators, 13

Turf, 185

Tuyeres, 215

SUBJECT INDEX

247

Unattackablb electrodes, 45
Uniform deposits, 44
Uranium, 138, 153. 236,
carbide, 153, 189

Vanadium, 24, 137, 153, 218, 236

carbon, 166
Vapour pressure of zinc, 141
Velocity of reaction, 9, 70
Viscosity of slags, 179, 221

Waste gas, 209

Water cooling, 179
Water gas, 197

power, 12
Whitewashing electrodes, 128
Wolframite, 233

Zinc, 8, 21, 25, 39, 58, 62, 84, 123,

139
Zinc scrap. 66, 71
Zircon, 25
Zirconium, 25, 153
Zones of reduction, 208

THE KND

BaiUiire, Tindall & Cox, 8, Henrietta Street, Covert Garden, W.C. 1

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