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

















t BY 

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











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 


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 


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. 



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 

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



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. 

June, 19 1 S. 






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 



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 



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

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


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

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





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



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


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

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


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


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 


INDEX . 239 



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. 





OH' 5-6 

K 2 05 

cr 2-i2 

Na i'i6 

NO's i'9i 

Ag i*66 




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 

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 


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 


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 


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 -^ 

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 


fa— fx—Vve 

But fa^+RT log aC 

hence Vas _ft-ri+RT logaG 

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 


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

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. 


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 

In many cases the temperature coefficient - is so small 

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


lations of n. 


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. 


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 



167 aOH'->H f O+0, 

' Ih-»h 2 

S0 4 "-»H 2 S0 4 +0 2 

1 "95 

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 


Caspari, the following are the values of the over- 

potential i\ : — 


1} in volts. 


i '3 








I '00 






Pt (bright) 


Pt (black) 


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. 


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 : 


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. 


* Electrolysis in Fused Electrolytes and Electro- 


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 


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. 


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 


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 

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. 


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 


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 


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 


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. 


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. 

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 

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- 

* Norwegian Royal Commission, Sept. 191 5. 


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- 

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 


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 


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 
Sulphuric acid 
Pan coke and breeze 


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. 


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. 


The Metaujferous Resources of the British 


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 


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 

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

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 

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 

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

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


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

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 : — 



Metric tons. 

United Kingdom 

. . 1915 



. . 1914 

. . 693,824 


. . I915 



• • 1915 


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 



Molybdenum is also found in England, Scotland and 

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 : — 


Metric tons. 

United Kingdom 


Malay States 

. . 329 

.. 2,326 





.. 663 
.. 83 

Victoria (1913) 

New Zealand (1913) 

Tasmania (1913) 


.. .. 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 


vanadinite, whilst smaller quantities are found in Western 

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 

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 


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. 


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. 



" 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. 




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 

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 


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 

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 

(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 


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 


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 

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, 



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 

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 















Al 2 O s 


Na 2 


K 8 


S0 8 

. 12275 

a ..' 


Free acid as H 2 S0 4 


Solids on ignition 

. 189/40 




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.). 

(c) The Sulphate Process, using Sulphur Dioxide as a 

Occasionally sulphur dioxide is injected into the anode 

I*. 3 


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 

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, 


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 : — 

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, 


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 

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 


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. 



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 




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 


to still further heat it to about 55 £ with exhaust 

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 . . 












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 



















OO3 2 










30 oz. per ton 



^q oz. per ton 

1 '17 


■ — 








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 

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 


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 



1 Alloys 

LSbo C 













*• 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. 



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 

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 


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. 


Watt recommends the following solution : — 

Copper sulphate 

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 . . 

> . . . 
• . . . 

. . . 

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


Water . . ■, . 

. . . 
. . . 

10 gms. 
10 litres. 

(b) NaHS0 3 •• 
Na 2 C0 3 

. . . 

• . * 

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 

. . . 
. . . 

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, 



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 

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 

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 


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 

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 



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 


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 

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. 


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 

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» 



or t — ~w 




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. 


0*0635 mm * 


0*0565 „ 


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 


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 


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 

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, 

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 

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. 


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

Cu(CN)' 2 ^Ctt+2(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 


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 


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 


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

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 

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 — 


— 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 : — 


—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 



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). 


(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 : — 


(1) Copper sulphate 25 gms. 

Zinc sulphate 29 gms. 

Potassium cyanide to dissolve the precipitate. 

• • 


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


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

. . 

. . 

. * 

• • 

• • 

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 

(3) Copper sulphate 
Stannic chloride 
Potassium hydroxide 

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

• • 

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. 



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 


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. 

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 


(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 


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 

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 


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 : — 



Lead peroxide 

Platinum (black) 

Platinum (bright) 


Volts (overvoltagc 
to oxygen). 

. 0-05 
. 0-24 
. 0-28 
. 024 
. 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 


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 




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 


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 

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 

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 


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 



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 


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 — 


> » 

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 


employment of sulphate electrolytes, more than the chloride 

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 

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. 


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 

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 


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. 


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 


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. 


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 


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. 


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 

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 — 


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. 


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 


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 


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 


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 


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 


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 

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 — 


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 


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- 



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. 


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 

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 


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. 



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 

Cyanide Electrolytes. — The cyanide complex silver 

Iy. 6 


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 

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 


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 

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 


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 


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 



• •' < 

. 0005 

Ag . 

• ■ < 

. 0*141 


■ • < 



• • < 

. 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 


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 


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 

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 


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. .. .. 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. 


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. 


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. „ 


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. 



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 : — 

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 


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— 

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. 


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 


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. 


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 


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 — 


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



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- 


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


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 


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 

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 

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 


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 


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. 


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 


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 


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 — 


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 


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- 
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. 


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 

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 

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> 


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, 


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 

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 

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


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- 

(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 

(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 

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 


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. 


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. 
18 Zeit. Elektrochem., 9, 1903, 762. 

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. 

17 Zeit. Phys. Chem., p. 158; 1895. 

18 Siemens and Halske, Dingl. Pol., 258, 288 ; 1893. 
18 J.S.C., p. 710, vol. 105 ; 1914. 

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

81 Trans. Amer. Electrochem. Soc, 27, pp. 291 ; 191 4. 
88 Zeit. Elektrochem., 10, 688; 1904. 

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, 


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

88 Michailenko, /. Russ. Phys. Chem. Soc, 44, 507; 191 2. 
88 Trans. Amer. Electrochem. Soc, 19, 137; 1911. 

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. 
84 " Elektroplattierung." Wien, 1900. 
88 /. Pract. Chem., 28, 149. 

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. 
48 See Met. Ind., 9, 479, and 8, 464. 
44 Met. Ind., 10, 384. 
46 Kern, Met. Chem. Eng., 9, 443. 

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

C. R. 145, 45. 

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. 


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 ; 

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. 


"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. 





" 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, 

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


Transactions of the Electrochemical Society. 

Transactions of the Faraday Society. 

Electrochemist and Metallurgist. 


L' Industrie Electrochemique. 


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. 





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


Gas rings Co eommtnot 


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 


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 


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 


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 


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 


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 












»• lo lo 4t ib to 70 flO so 



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 




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 

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

(C) Processes using a molten lead cathode. 

(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 

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 


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 

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 


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 


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 


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


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. 


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 


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 

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 


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. 


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 

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 


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. 


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- 

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 


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 


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. 


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 


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 


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. 


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 

The composition of bauxite varies with the source ; 
the following represent typical analyses : — 

Per cent. 





A1 2 3 

• • 56 



54' 1 

Fe 2 3 

•• 3 




0.02 • • 

. . 12 




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 


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 


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 

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. 



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. 


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 ( 



\ / 


\ / 


r \ / 


1 f 





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 
















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 

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 


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 


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 


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 : — 


Carbon electrodes 
Labour, repairs, interest 

on capital, superin- 


Costs per kgm. 







4 - 4 

Carbon electrodes . . 










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. 


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. 



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. 


" 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. 



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). 


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. 


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 

thallium salts in terms of the electrolytic potential difference 

Tryrr is— 


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 . 
Nd , 
Pr . 

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. 










CeCl 2 


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 


required to ensure the fusion of the metal. The addition 
of J part by weight of barium chloride to the chloride is 


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). 


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. 


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. 



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. 


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 


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. 


" Die Darstellung des Chromes und Seiner Verbindungen." W. Le 
Blanc. Halle. 1902. 



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 

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." 


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 


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. 






















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 


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 

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 


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 


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 



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 . . 









0'3 oz 


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 


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 — 


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 



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 

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 

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. 





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) 


can be smelted, directly separating the metal from the 

(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 

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 : — 




Cu .. 

- 510 



Fe .. 

... 2850 




. . 7-64 



S .. 

4 - I2 



Al 2 O s 





-. 7*30 


Si0 2 

. . 2370 


The power consumption for the furnaces averaged 4750 amps. 
at 119 volts with a power factor of 0*9. 


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. 


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 



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 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.). 


Harden 3° has given details of the conditions necessary 
for the reduction of tin ores. Although electrothermal 


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. 


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 



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 



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. 


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. 


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 


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. 


2 24 

1 96 





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- 

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). 


Carbon can exist in at least three well-known modifica- 
tions, two crystalline 47 and one amorphous : diamond, 


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. . 


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 

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. 


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. 




Processes of electrical graphitization have been applied to 
various grades of coals, and even to dried peat, but have not 
proved technically successful. 


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 


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 

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. 


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. 


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. 



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 


-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. 


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. 


• " 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. 



"The Electric Furnace," Stansfield. 

"Der Elektrische Ofen," Bronn. 

" Kunstlicher Graphit," F. A. Fitzgerald. 1904. 

"Die Metallurgie des Zinns," H. Henniche« 



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 


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 

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. — 


Although the explanation of the production of siloxicon 


(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 
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. 







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 



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 



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. 


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 


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. 


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. 


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 


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. 


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/' F. A. J. Fitzgerald. 1904. 
"The Electric Furnace," Stansfield. 
■'Applied Electrochemistry," A. Allmand. 


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 



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 

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 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 

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- 

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. . . 





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 


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 

CaC 2 = 3,900 

per gram 

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 


Net energy required=i66,700 calories or 

= 3,040 kw. hours per metric 


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- 

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 


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 


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 


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 

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 


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. 


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 

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 


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. 


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. 



"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. 



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 



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. 


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 

C. The fixation of atmospheric nitrogen by biochemical 

D. The preparation of combined nitrogen from which 


ammonia can be obtained, e.g. cyanamides, cyanides and 

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 


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- 

(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 


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 


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 : — 


Percentage NH 8 







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 


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 


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 

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 


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 

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. 


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 


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 


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. 

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- 

The Cyanides. — Since the market value of fixed 


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. 


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. 


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. 


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 


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 


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 


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 


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 


ton nitrogen. 

metric ton N 2 . 

of energy. 





Hatisser . . 



Gas, 30,000 
cu. m. Cal. 
value 4300 
per cu. m. 
=3 150,000 
kw. hrs. or 
kgm. coal. 

Haber . . 

For gas com- 

For prep. 




ofH 2 


For N 2 pre- 


For heating ca- 

Water gas 



production 2,400 

For circulation 


H 2 

of gases 



on 1,600 


4,000 kgm. 

Serpek . . 

For N 2 pre- 




For reduction 

and azotising 



Bucher . . 

For N 2 pre- 





For production 

of the carbide 



Forazotizing . . 


For grinding . . 


For production 

of the nitrogen 



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 


either the Bucher or modified Serpek (see p. 1Q2) processes 
would be serious rivals to the Haber or cyanamide. 10 


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. 


"The Cyanide Industry," R. Robine & M. Lenglen. Wiley & Sons. 

"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. 


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) 


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. 


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 — 


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 


about 2:1. The waste gas is generally used for steam 

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 

The following analyses of iron ores represent those used 
in actual practice : — 




Fe 2 3 . 

• 73-840 



Fe s 4 . 



MnO . 

• 0567 



Si0 8 . 




A1 2 3 . 








MgO . 


I 030 









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 


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 

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 

. . - • 


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. 



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 

Furnaces. — The earliest successful furnaces employed 


for the production of pig iron were those of Keller 
and Heroult, originally used for the production of ferro- 

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> 


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 

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 


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 « 


ia. 22. — Shaft furnace for 


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 

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. 


(b) Fusion of pure materials. 

(c) Fusion of pig iron, scrap steel with fluxes, with or 
without the addition of oxide of iron. 

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 


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 


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 — 


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 — 



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 : — 


Al.O,. CaO. 

MgO. CaS. 


Softening tem- 
perature °C. 







.. 1244 












. . 1279 
. . 1262 
. . 1263 


Per cent. 


Softening tern 







peratuie C 






22 . 

• 1297 






07 . 

• 1331 






0*5 • 

• 1352 






0-3 . 

• 1342 






o*6 . 

• 1343 






0*5 • 

• 1358 






03 . 

• 1365 






O'l . 

• 1356 






o\5 • 

• 1383 






o'5 . 

• 1425 






0'2 . 

• 1403 






0*2 . 

. 1388 






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. 


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. 


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-«. 


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 


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. 




. . 0*80 



.. 0*30 



. . 0*22 


Phosphorus . . 

.. 0*015 



.. 0-045 




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 


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 


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 



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 





////// /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 


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 


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 

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 


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 

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. 


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. 


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. 



Tungsten . . 

.. 85-15 



. . 14*12 



. . 0-45 



. . 0-13 


Manganese . . 

. . 0-085 



. . 0'02I 



. . o # oi8 



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 

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- 

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, 


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. 


.. 2705 


.. 425 


. . o-6o 


. . 046 


. . 0*22 


. . 0-3I 


. . 0*02 


. . 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 : — 



Percentage composition; 


. . 80*20 









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 

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 


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. 


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". 


" 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." 







Calcium . . . 




Cadmium ... 

Thallium ... 



Hydrogen . . . 


Bismuth ... 
Mercury ... 
Platinum ... 


Potential E k . 

lytic over- 
q to H a . 




+ 2*56 

+ ? 

+ 1-49 

+ ? 
+ ? * 
+ 1*075 


(in Tl so- 









lent wt. 
of metal 
in grras. 
per am- 
















1 63 






















|* # 45 


Usual current densities employed for 

deposition, refining, plating in amperes 

per sq. decimetre. 











{Castner 200 

110-250, con- 
tact elec- 
trode up to 
/Hall 100 1 
\Heroult 190] 


Chloride | 

electro- 73-4 
1 lyte ) 

1 -3-1-5 






















1-2 O'H 




ride 2-3 J 





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 




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 

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 



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 


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 



Watts, 65, 103, 104, 207 
Weidlein, 34 
Weintraub, 170 
Westman, 159 
Whetham, 2 
Whitney, 47, 48 
Wilson, 177 


Winogradsky, 191 
Withrow, 76 
W6hler, 120 
Wohlwill, 79 
Wood, 136 
Woolrich, 69 
Wright, 132 


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, 

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 



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, 

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 

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 



Gelatine, 75, 80, 84, 87, 103 

Glucosides, 103 

Glue, 83, 87, 103 

Glycerine, 113 

Gold. 39, 69, 84, 71, 74, 75, 79, 84, 

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, 

Hexaminetetramine, 206 
Homogeneous reactions, 217, 219, 

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 


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, 

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 




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 


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 
nitride, 195 

Transmission of power, 15 

Transport numbers, 1 

Tungsten, 8, 153, 187, 218, 234 
carbide, 233 

Turbo-generators, 13 

Turf, 185 

Tuyeres, 215 



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, 

Zinc scrap. 66, 71 
Zircon, 25 
Zirconium, 25, 153 
Zones of reduction, 208 


BaiUiire, Tindall & Cox, 8, Henrietta Street, Covert Garden, W.C. 1