Handbook on Rare Earth Metals and Alloys (Properties, Extraction, Preparation and Applications)

Handbook on Rare Earth Metals and Alloys (Properties, Extraction, Preparation and Applications)

Author: NPCS Board of Consultants & Engineers
Format: Paperback
ISBN: 9788178331201
Code: NI218
Pages: 688
Price: Rs. 1,875.00   US$ 150.00

Published: 2009
Publisher: Asia Pacific Business Press Inc.
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Rare earths are essential constituents of more than 100 mineral species and present in many more through substitution. They have a marked geochemical affinity for calcium, titanium, niobium, zirconium, fluoride, phosphate and carbonate ions. Industrially important minerals, which are utilized at present for rare earths production, are essentially three, namely monazite, bastnasite and xenotime. In modern time techniques for exploration of rare earths and yttrium minerals include geologic identification of environments of deposition and surface as well as airborne reconnaissance with magnetometric and radiometric equipment. There are numerous applications of rare earths such as in glass making industry, cracking catalysts, electronic and optoelectronic devices, medical technology, nuclear technology, agriculture, plastic industry etc. Lot of metals and alloys called rare earth are lying in the earth which required to be processed. Some of the important elements extracted from rare earths are uranium, lithium, beryllium, selenium, platinum metals, tantalum, silicon, molybdenum, manganese, chromium, cadmium, titanium, tungsten, zirconium etc. There are different methods involved in production of metals and non metals from rare earths for example; separation, primary crushing, secondary crushing, wet grinding, dry grinding etc. The rare earths are silver, silverymwhite, or gray metals; they have a high luster, but tarnish readily in air, have high electrical conductivity. The rare earths share many common properties this makes them difficult to separate or even distinguish from each other. There are very small differences in solubility and complex formation between the rare earths. The rare earth metals naturally occur together in minerals. Rare earths are found with non metals, usually in the 3+ oxidation state. At present all the rare earth resources in India are in the form of placer monazite deposits, which also carry other industrially important minerals like ilmenite, rutile, zircon, sillimanite and garnet.
  Some of the fundamentals of the book are commercially important rare earth minerals,   exploration for rare earth resources, rare earth resources of the world, some rare earth minerals and their approximate compositions, rare earths in cracking catalysts, rare earth based phosphors, interdependence of applications and production of rare earths, uranium alloys, conversion of ores to lithium chemicals, characterization and analysis of very pure silicon, derivation of molybdenum metal, electoplating and chromizing, electrolytic production of titanium, heat treatment of titanium alloys, tensile properties of alloys etc.
The book covers occurrence of rare earth, resources of the world, production of lithium metals, compounds derived from the metals, chemical properties of beryllium, uses of selenium, derivation of molybdenum metals, ore concentration and treatment and many more. This is a unique book of its kind, which will be a great asset for scientists, researchers, technocrats and entrepreneurs.

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Natural Abundance
Occurrence of Rare Earths
1. Placer Deposits
2. Vein Type Deposits
3. Bastnasite Deposits
4. Ion Adsorption Type Ores
5. Other Rare Earth Sources
Commercially important Rare Earth Minerals
1. Monazite
2. Bastnasite
3. Xenotime
Exploration for Rare Earth Resources
Rare Earth Resources of the world
1. China
2. United States of America
3. India
1. Beach Placers and Dunes
2. Inland Placers
3. Reserves of Monazite
4. Occurrence of Xenotime
4. Australia
1. Placer Deposits
2. Hard rock Deposits
1. Mount Weld Deposit
2. Mary Kathleen
3. Port Pirie
4. Olympic Dam
5. Brazil
Other countries
Table 1.
The Content of Rare Earths and Some Common Elements in the lgneous Rocks of the Earth's CrustTable 2.
Some Rare Earth Minerals and their Approximate Compositions
Table 3.
Typical Placer Minerals and their Specific Gravity
Table 4.
Mineralogial Composition of Typical Placer Samples, as mined in India and Australia
Table 5.
Rare Earth Distribution in Various Rock Forming and Accessory Mineral of Host Rock
Table 6.
The Rrare Earth Pattern in Different Layers of an Ion Adsorption Type Desposit
Table 7.
Composition of REO recovered from major Ion-Adsorption Type Deposits in China
Table 8.
The REO Centent of Different Types of Ores in China
Table 9.
Analysis of the Typical Loparite Sample
Table 10.
Rare Earths Distribution in Monazite from Different Sources (wt %)
Table 11.
Rere Earth Distribution in REO from Bastnasite from different
      Sources (wt. %)
Table 12.
Rare Earth Distribution in Xenotime Samples (wt.%)
Table 13.
Ore Types In Baiyunebo Deposit
Table 14.
Chemical Analysis of Ore Samples from Deposit no. 801, China
Table 15
Some Important Rare Earth Resources of Australia and their Rare Earth and Thorium content
Table 16
Countrywise Distribution of Rare Earth Resources
1. Arc Carbons
Glass Making Industry
1. Decolourization of glass
2. Colouring of glass
3. Special Glasses
1. Spectacle Glass
2. Television and Cathode Ray Tubes
3. Glass for Eye protection
4. Infrared Transmitting Glass
5. Radiation Protection Windows
6. Optical Glass
Laser Glass
Glass Polishing Powders
1. Glass Polishing Technology
2. Different Types of Abrasives
3. Manufacturing Methods
4. Producers of Polishing Powders
4. Enamels and Glazes
1. Rare Earths in Cracking Catalysts .
1. Cracking Process
2. Evolution of the catalyst
3. Rare Earth, Exchange of the Zeolite
4. Composition of the catalyst
5. Role of Rare Earths in the Catalyst
Use of Rare Earth Zeolites
6. Rare Earth Consumption
7. Impact of Lead Additive Phase down
8. Scope for using cerium in FCC unit
2. Application of Cerium and Lanthanum in Auto-exhaust Catalysts
1. Catalyst Converter System
2. Role of Rare Earths
3. Other Catalyst Applications of Rare Earths
1. Methanation
2. Ammonia Synthesis
3. Homogenous Catalysis
4. Methane Conversion
Fine Ceramics
1. HighTemperature Structural Ceramics
1. Stabilization of Zirconia
2. Sintering of Silicon Nitride (Si3N4)
3. Sintering of Silicon Carbide (SiC)
2. Functional Ceramics
1. Piezoelectric Materials
1. Role of REO in Piezoelectric Ceramics
2. Applications of Piezoelectric Ceramics
2. Optoelectronic Materials
1. Applications
2. Preparation of PLZT Materials
3. Thermistor, Varistor and Capacitor Materials
1. PTC Thermistor
2. Varistor Materials
3. Grain Boundary Barrier Layer (GBBL) Capacitors
4. Solid Oxide Fuel Cells
1. Electrolyte
2. Electrodes
3. Interconnecting Material
5. Oxygen Sensors
6. Heating Elements
7. High Temperature Super-conducting Materials
Rare Earth Based Phosphors
1. General
1. Laser Action
2. Antistoke Emission
2. Rare Earths as Phosphor Materials
1. Fluorescence due to 4f Transitions
2. Fluorescence due to Transitions from 5d to 4f Orbital
3. Rare Earths as Phosphor Matrices
3. Major Applications of Rare Earth Phosphors
1. Low Pressure Mercury Lamps
1. Desirable Phosphor Properties for Fluorescent Tubes
2. Phosphors used in Tube Lights
3. Rare Earth Phosphors in Fluorescent Tubes
2. Rare Earths in High Pressure Mercury Vapour Lamps
3. Trichromatic Compact Lamps
1. Matching of Lamp Light to the Visual System
2. Red Phosphor
3. Green Phosphor
4. Blue Phosphor
5. Performance of the Trichromatic Lamp
4. R&D in phosphor Development in India
5. Preparation of Light Phosphors
6. Application of Cathodoluminescence of Rare Earth
1. Colour Television Phosphors
2. Preparation of Phosphors
7. Phosphors for Non-illumination Purposes
8. Electroluminescent Phosphors
9. Thermoluminescent Phosphors
10. Rare Earth X-ray phosphors
1. X-ray screens and scanners
2. Advantage of Rare Earth Phosphors
3. Rare Earth Compounds used in X-ray phosphors
11. Rare Earths in other Medical Imagery
Rare Earths in Nuclear Technology
9. Miscellaneous Applications
1. Application in Agriculture
1. Techniques of Application
2. Nong-le and N.P.K. Fertilizers
3. Areas of Application
2. Dyeing and Currying
3. Colouring of Plastics
Interdependence of Applications and Production of Rare Earths
Particle Characteristics
Table 1.
Types of Middling
Staged Concentration
Gravity Separation
Chemical Methods
Magnetic and Electric Methods
Exploitable Factors
Concentration Formulae
Crushing Theory
Physical Aspects of Comminution
The Crushing Sequence
Jaw Crushers
Variations on the Blake
The Dodge Crusher
Gyratory Crushers
Comparison of Jaw and Gyratory Crushers
Mobile Crushing Units
By-passing the Undersize
Feeding Arrangements
Protective Devices
The Duty of the Section
Lay-out and Equipment
The Symons Cone Crusher
Gearless Gyratories
Hammer Mills
Gravity Stamps
Dry Crushers, Summarised
Optimum Grind
Applied Power
Useful or Net Power
Grinding and the Particle
Grinding Objectives
Comminution of Particles
Effect of Peripheral Speed
The Return Load
The Solid-Liquid Ratio
Fixed-path Mills
The Vibrating Mill
Tumbling Mills
Mill Capacity
General Conclusions
Milling Action
Types of Mill
The Hardinge Mill
The Low-discharge Cylindrical Mill
Tube, or High-discharge Mills
The Cascade Mill
Mill Liners
Crushing Bodies
Isotopes and Nuclear Reactions
Elastic Properties
Tensile Properties
Deformation and Textures
Recovery, Recrystallization, and Grain Growth
Reactions with Nonmetallic Elements; Binary Compounds
Reactions with Simple Compounds of Nonmetallic Elements
Reactions with Aqueous Solutions
Uranium Alloys
Nonmetals: Carbon, Boron, and Silicon
Liquid Metals
Phase Diagrams
Table 13. Alloying Behavior of Uranium
Melting and Casting
Swaging and Drawing
Powder Metallurgy
In Nuclear Reactors
Other Uses
Cost Considerations
Production of Lithium Metal by Fused Salt Electrolysis
Lithium Cartridges
Lithium Wire or Ribbon
Lithium Shot
Sodium-Free Lithium Metal
Molten Lithium
Lithium and Hydrogen
Lithium and Nitrogen
Lithium and Oxygen
Lithium and Silicon
Lithium Hydroxide
Lithium Halides
Various Other Lithium Compounds
Lithium-Magnesium Alloys
Lithium-Aluminium Alloys
Lithium-Zinc Alloys
Lithium-Lead Alloys
Lithium in Alloys
Lithium as a Degasifier and Refining Agent
Lithium in Cast Iron
Lithium in Steels
Lithium in Organic Chemistry
Lithium in Atomic-Energy Developments
Lithium in High-Energy Fuels
Copeaux-Kawecki Process
Sawyer-Kjellgren Process
Pure Beryllium Oxide
Beryllium Metal
Beryllium-Copper Master Alloy
Beryllium Oxide
Beryllium Alloys
Beryllium-Copper Alloys
Beryllium Oxide
Beryllium-Copper Alloys
Beryllium-Nickel Alloys
Beryllium-Iron Alloys
Miscellaneous Beryllium Alloys
The Solid State
The Liquid State
The Vapour State
Electrical Conductivity
Effect of Light on Electrical Properties of Selenium
Electronics Industry
Glass and Ceramics Industry
Pigment Industry
Steel Industry
Miscellaneous Uses
Extraction of Platinum Metals from
Canadian Nickel Ores
Extraction of Platinum from South African Ores
Refining of Platinum Metal Concentrates
Treatment of Native Platinum
Refining of Scrap
Vapour Deposition
Available Forms
Rhodium and Iridium
Ruthenium and Osmium
Alloys of the Platinum Metals
Compact Metals
Sponge and Powdered Metals
“Blacks” and Colloidal Metals
Occurrence and Sources
Production and Price Statistics
Production of Tantalum Metal
Consolidation and Purification
Physical Properties
Mechanical Properties
Chemical Properties
Tantalum-Tungsten Alloys
Nuclear Energy Systems
Physical Properties
Mechanical Properties
Calcium Hydride
Calcium Alloys
Powder Metallurgy Process
Arc-Casting Process
Corrosion Resistance of Metallic Molybdenum
Molybdenum as an Alloying Element
Alumino- and Silicothermic Chromium
Carbon-Reduced Chromium
Electrolytic Chromium
Electronic Structure
Thermal Properties
Electoplating and Chromizing
Initial Recovery
Binary Systems
Ternary Systems
Oxide Reduction
Magnesium Reduction of Titanium Tetrachloride
Sodium Reduction Titanium Terachloride
The lodide Process
Electrolytic Production of Titanium
Chemical Compounds
Alloying Principles
Heat Treatment of Titanium Alloys
Tungsten Compounds
Tungsten Metal
Tungsten Carbide
Tungsten-Molybdenum, Columbium, Tantalum, Chromium
Tungsten Steels
Ductile Rod and Wire
Tungsten Sheet
Slip Casting
Arc Casting
Electron Beam Melting
Hydrostatic Compacting
Flame Spraying
Sintered Carbide
Present Applications
Potential Applications
Sources of zirconium
Separation of Zirconium and Hafnium
Reduction of ZrO2
Reduction of Zirconium Halides
Reduction of Other Compounds
Reduction of ZrCI1 with Mg—the Kroll Process
Iodide Decomposition Process
Electrodeposition of Zirconium
Physical and mechanical properties
Reaction with Gases
Reaction with Halogens
Corrosion in Various Media
Corrosion in Gases
Corrosion in Liquid Metals
Corrosion in Other Media
Rolling and Forging
Cold Working
Power Brake Forming
Surface Finishing
Tensile Properties of Alloys

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Sample Chapters

(Following is an extract of the content from the book)
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The group of 15 chemically similar elements with atomic numbers 57 through 71 inclusive and yttrium (atomic number 39) are considered Rare Earths. Though two misconceptions concerning rare earths stem from the name itself  such as  they  are rare  and they  are earths  it needs to be stressed that both are only historical hangovers and they are  in fact  neither rare nor earths. This realization is essential to appreciate the role they play in modern life and technology and to not only explore but also to achieve new and rewarding applications for them.

Natural Abundance

All the rare earths except promethium (Z 61) occur on the planet earth. The absence of promethium in terrestrial sources is due to its radioactive nature coupled with a short half life (t½) of only 2.7 years. Rare earths are widely dispersed in many minerals. For knowing the relative abundance of individual rare earths in the earth as a whole  it is useful to consider the composition of a group of meteorites called carbonaceous chondrites. There are several geochemical evidences to show that the earth and these meteorites had a common origin. The data of Schmitt et al1 on the rare earth content of these meteorites shown in Fig. 1 indicate that the abundance decreases with increasing atomic number and significantly a rare earth with even atomic number is more abundant than its odd atomic number neighbours. Elements are enriched or depleted in the igneous rocks relative to chondrites and for this a number of geochemical reasons have been advanced. The concentration of rare earths in the igneous rocks of the crust is much higher relative to that of the earth  as a whole. The crustal abundance data for rare earths and for some common metals show that at least some of the former are no less abundant  if not more abundant than some of the latter.

Occurrence of Rare Earths

Rare earths are essential constituents of more than 100 mineral species and are present in many more through substitution. They have a marked geochemical affinity for calcium  titanium  niobium  zirconium  fluoride  phosphate and carbonate ions. Based on chemical type nine groups of rare earth minerals can be distinguished. They are  (1) fluorides  (2) carbonates and fluoro carbonates  (3) phosphates  (4) silicates  (5) oxides  (6) arsenates  (7) borates  (8) sulphates and (9) vanadates. The first five of these are of significance for the recovery of rare earths  industrially. Some typical examples of each of these types and their approximate chemical composition are shown in Table 2.

The similar atomic structure of rare earth elements is reflected not only in their chemical and physical properties but also in their close geochemical association in nature. The rare earth bearing minerals carry a suite of all the elements of this group. However  the relative abundance of individual members is not the same in all minerals. There is a preferential enrichment of either the cerium group (light rare earths) or the yttrium group (heavy rare earths) among the minerals. Monazite  bastnasite and other minerals bearing light rare earths are more abundant than minerals like xenotime  which carry the heavier members of the family. Industrial production of rare earths  at present  is confined essentially to the three minerals   monazite and bastnasite  which are rich in light rare earths and xenotime  which is rich in yttrium and heavy rare earths. The geological environments in which the rare earth minerals are found vary widely.

Placer Deposits

Placer deposits are the major source of monazite and xenotime  the two commercially important phosphate minerals. Placer minerals are formed by natural erosion and weathering of igneous and metamorphic rocks in which the two rare earth minerals along with ilmenite  rutile  zircon  magnetite and a few others occur in small concentration. Natural agencies like water  ice  atmospheric gases and sun play an important role in these processes. Streams and rivers transport the weathered part of the rock and carry the released mineral particles. During their transportation two forces act on the particulate matter  (1) the current  which carries them forward and (2) their own weight  which pulls them down. The eddy currents lift some of the particles that settle and push them forward. While clay minerals get finer and finer as they travel down stream  hard  heavy and chemically resistant particles get abraded and become smooth and round. As the velocity of the current decreases  some of the heavy minerals get concentrated in traps within the river system itself. These are the alluvial placers. Some of the heavy minerals reach the sea.

Though the mineral sands are expected to concentrate near and around the river mouths  as the velocity of flow decreases considerably at that point  the waves approaching the shore at an angle distribute them along the coast. Therefore  placer deposits usually extend over several kilometers  on either side  from the mouth of the river. The breaking waves pick up all available loose sand of the near shore material and throw it on the beach. On the other hand  the receding water  which has lost most of its kinetic energy carries back to the sea  preferentially the light material. If the process of sorting proceeds for a long enough time it leads to the formation of beach placers. The nature and richness of placer deposits depend as much on the conditions of the sea and the beach as that of the source rocks. Beach placers usually occur in the narrow strip between the high and low tide levels. Concentration of heavy minerals in the bottom layers of the sand in the shallow waters near the coast is also common. During the ice ages  when the sea level was much lower than at present  several rivers would have had to flow further than at present to meet the sea. Deposition of placer minerals would have taken place in their valleys  but at present thick sediments have covered them. If they are not at too great a depth  they can be mined from under the waters.

Some of the typical minerals  which are associated with monazite and xenotime in the placer deposits  are shown in Table 3.

Monazite and xenotime are usually minor constituents in the placers and they are recovered as byproducts  but in the form of high grade concentrates  while undertaking the commercial production of other heavy minerals like ilmenite  rutile  zircon and cassiterite.

Large beach  coastal and inland placer deposits occur in many places spanning the continents of Africa  Asia  America and Australia. The important ones are the deposits in Australia  India  Brazil  South Africa  USA and China. Similar deposits  which contain a combination of monazite and xenotime are found in the form of tin bearing alluvial placers in Malaysia and other south east Asian countries. In case of placer deposits nature has conveniently concentrated the heavy minerals  thus eliminating the need for extensive drilling   blasting and  in some cases  acid leaching. It can  generally  be said that placer resources of rare earths have more chance to come to fruition than their hard rock counterparts  when discovered.

Vein Type Deposits

Stray occurrence of rich veins having a high content of monazite has been found in some parts of the world. One such deposit was discovered in 1950 at Steen Kampskrall in the Republic of South Africa. At that time it was considered to be the richest source of monazite. The other primary monazite deposits of commercial importance are located in Colorado (USA) and Baiyunebo (China).

Bastnasite Deposits

Bastnasite occurs in a finely divided form or as pheno cryst in carbonatites. Quartz veins and epithermal fluorite bearing veins also contain this mineral. The first major discovery of a bastnasite deposit was at Mountain Pass in California in 1949. For quite some time after its discovery this was considered the world s largest rare earth deposit. Bastnasite occurs here in the carbonatite bodies intruding a precambrian basement complex. The main orebody  mined by MolyCorp is about 750 m long and up to 210 m wide. The ore contains 12% bastnasite  25% barytes  10% quartz  10% strontianite and 40% carbonates.

Rare earth bearing carbonatites are also located in some parts of Africa  including Burundi  Kenya and Malawi. Some of these also yield pyrochlore and monazite. A carbonatite complex located m Minas Gerais region of Brazil can be a good source of niobium  monazite and a few rare minerals.

A rare earth deposit of great importance and presently the world s largest one is in Baiyunebo  Inner Mongolia  China. Here bastnasite and monazite occur together as associated minerals in a major iron ore deposit. Somewhat similar deposits are also identified in South West and South East parts of that country.

Ion Adsorption Type Ores

The ion adsorption type deposits are weathered residues where rare earths are adsorbed on clay type minerals. Commercially attractive quantities of this ore type have been found  so far  only in parts of China and they play a very significant role in the development of rare earth industry of that country. They have already attracted the interest of other countries as a source material. This material has some desirable characteristics  which no other type of ore possesses.

The ion adsorption deposits are the result of in situ weathering of rare earth rich host rocks  most commonly granitic and volcanic types. There are two important prerequisites for the formation of these deposits. First  there must be a sufficiently large body of rare earth bearing host rock  occurring within the zone of weathering and second  the weathering or lateritic process must be for a prolonged period with limited erosion. This requires a mild  rainy and techtonically stable environment over a geologically long period of time. In Southern China  where both conditions are met  a large number of deposits of this type have been found. The granitic rocks of Yinshen age (195 130 million years old) are the most common host rocks for such deposits. Bastnasite  synchisite  allanite and perrierise are the rare earth bearing minerals present in the weathered zone of these ion adsorption deposits. The stability of the rock forming minerals and accessory minerals like plagioclase  biotite  manganese garnet  titanite and epidote  to the chemical weathering process is also a critical factor in the formation of these deposits. In many formations of this type the rare earths are distributed in both the rock forming and accessory minerals of the weathering host rocks.

Weathering is the most important factor in the formation of ion adsorption deposits in which several horizons can be distinguished such as (i) an upper horizon of colluvium and soil underlain by (ii) a thick zone of strongly weathered rock enriched in rare earth elements (iii) a less weathered transitional zone over (iv) a basal hard rock.

One important feature  which distinguishes the ion adsorption ores of China from any other ore in the world  is the depletion of cerium  particularly in the strongly weathered zone of the deposits. Consequently  there is lanthanum enrichment in these materials. The highly oxidizing conditions prevalent in the upper leach regions of the deposit  result in oxidation of Ce3+ to Ce4+  leading to its hydrolysis and retention by the top soil layers. Middle rare earths like samarium and heavies are more abundant in the material from deeply weathered zones.

The characteristics of the ion adsorption type ores which make them attractive for rare earth producers are  (i) low cerium content  (ii) relatively high yttrium content in some and middle rare earths content in others and (iii) easy mining and processing to recover rare earths. Table 7. shows the varying composition of the rare earth oxide product recovered from different ores of this type  originating from different mines in China.

The actual rare earth content of the ion adsorption type of ores is much less than that of other types of ores. However  a major advantage with these ores is their amenability to a simple chemical leaching  which leads to a high grade rare earth oxide product with very few processing steps.

The highly variable rare earth composition of the ion adsorption ores with respect to combined lanthanum and cerium content as well as other rare earths content makes these ores very attractive for selective mining to meet the required production targets of individual rare earths. This aspect  more than the total rare earth reserves in this form plays an important role in China s strategy for producing individual rare earths.

Other Rare Earth Sources

In addition to the types of deposits mentioned  there are others in some countries  which play a role in providing rare earths in significant quantities.




Early applications of rare earths on an industrial scale started with the use of their compounds  particularly the oxides and fluorides. Use of cerium in the Welsbach gas mantle  though as a minor addition to thorium nitrate  falls in line with this trend. This was followed by their use in several fields and it is surprising that some of these applications continue to the present time. Some of the important areas where compounds of rare earths are used industrially in varying degrees of purity are outlined here.

Arc Carbons

Soon after Welsbach commercialized gas mantle production  mixed rare earths started accumulating at all the monazite processing plants. Production of mischmetal for lighter flints provided an outlet for this byproduct. Another application was developed by Electrodes Bergenlicht Co. of Germany  by way of manufacture of arc carbons during World War I. This was essentially a military application for search lights and was based on the observation that in a carbon are the rare earths emit a bright light. Initially attempts were made to incorporate rare earths in the carbon electrodes in different ways. One such way was to soak the rods in a solution of mixed rare earths  but the effect was not good. After several modifications it was finally found that incorporating the rare earth compounds in the core of one of the electrodes  used for striking the arc  brought about the expected result. In the early version of such an electrode a thin metal wire was incorporated in the rare earth mixture to make it a good conductor. The process of electrode preparation underwent a change later.

It is interesting to note that in Japan the industrial processing of rare earths began with the manufacture of rare earth fluoride  during World War II  for military applications  such as fabrication of powerful searchlights. Today Japan stands as one of the major users of rare earths for the very mundane to the most sophisticated applications and has a variety of production facilities for rare earth products.

Radiant arc is the light source for search lights and in the civilian field  for motion picture projectors. Radiation of reasonable brightness is produced by an arc struck between two carbon electrodes. In a steady arc the prevailing high temperature vapourises and ionizes the carbon. However  the light from a carbon arc is basically of the low intensity type. When the era of colour motion pictures came into being it was realized that the simple carbon are was not bright enough to project these pictures on a wide screen. The need for development of a much brighter source of light with good spectral quality arose. It was in this context that cored carbon electrodes were developed.

Compounds of rare earths  usually mixed rare earth oxide and fluoride with minor addition of other metal salts  are kneaded with pitch and packed as core in the positive carbon electrode. The shell of the electrode is the main conductor. As the arc is struck between the electrodes the core vapourises faster than the shell (made of carbon)  thus forming a cup on the face of the positive carbon  which is the main source of light. The rare earth compounds vapourise from the cup shaped crater and emit brilliant light as they provide abundance of spectral lines  favourable energy distribution  high light intensity and good colour balance. The use of mixed rare earths is essential for getting balanced light.

It is reported that due to difference in composition of rare earths  the mixed rare earths product derived from bastnasite is not so good for this application as the one derived from monazite. Considerable quantity of mixed rare earth compounds is utilized for this purpose in the world. Up to 1960 s nearly one fourth of the rare earths produced in the world were used in this field. However  with the emergence of halogen lamp as radiant source  the demand for arc carbons has diminished to some extent.

Glass Making Industry

The applications of rare earths in the glass industry were among the very early ones to come into practice. Like in other fields in which rare earths are used  the scope of application is modified and expanded over the years. Starting with the use of naturally occurring mixtures there has been a gradual sophistication in use  needing moderate to high purity starting materials. The field of application can be conveniently split into several sectors  (1) decolourization of glass  (2) colouring of glass  (3) making of special glasses and (4) glass polishing.

Decolourization of glass

In the production of transparent glass presence of traces of divalent iron in the raw material is considered harmful as it imparts a bluish green colour to the manufactured glass. Oxidizing agents like manganese dioxide (MnO2) and arseneous oxide (As2O3) were used for a long time to reduce or eliminate such colour. In 1896 Dressbak patented and manufactured a mixture containing cerium and other rare earth oxides for decolourizing glass. This was  perhaps  the first large scale use of cerium. The beneficial effect of cerium is attributed to Ce4+ which being a strong oxidizing agent converts Fe2+in the glass melt to Fe3+ leaving only a faint yellow colour in the glass. The colour of the small amount of neodymium and praseodymium in the added rare earth mixture masks the residual yellow colour and makes the glass look colourless and transparent.

For the refining of glass by elimination of bubbles  cerium oxide is as good as other additives like arseneous oxide. However  the latter is a health hazard on account of its poisonous and carcinogenic nature.

When cerium is used to decolourize flint glass containers it absorbs ultraviolet light and thereby reduces the deterioration of ultra violet light sensitive materials like food  chemicals or medicines stored in such containers.

Colouring of glass

Rare earths are also well known as excellent colourants of glass. They can be used to manufacture both normal coloured glass and coloured optical glass.

Addition of 3.5% CeO2 to glass imparts beautiful yellow colour mixed with some light red. Addition of a mixture of CeO2 and titanium dioxide (TiO2) imparts to glass a bright golden yellow appearance.

Neodymium glass exhibits a beautiful wine red colour with soft tone and a remarkable double colour effect. Addition of Nd2O3 along with manganous oxide and selenium makes the glass take a lilac colour. Different colours show up in different lights.

Glass with praseodymium addition shows green colour in the sun and nearly colourless in candle light. Holmium and erbium impart light shades of colour but they are relatively costly for general use.

Rare earth colourants have been widely used in glassware production in China to make ornaments with beautiful bright colours2. The jewelry glass artificial crystals diamond pearl and agate made with rare earths have nearly 30 colours and tones. The application of rare earths in this field has good prospects.

Special Glasses

Apart from colouring and decolourizing of common forms of glass some of the rare earths proved useful for manufacture of glass for special purposes.

Spectacle Glass

In 1912 Crooks found that addition of cerium to ophthalmic glass makes the latter absorb ultraviolet light  a feature useful for eye protection. For this purpose high purity (+99.95%) cerium is used. For absorbing effectively all radiation below 3200A0 both CeO2 and TiO2 are added to glass. Praseodymium and neodymium are added as didymium oxide for colouring Crooks glass.

Television and Cathode Ray Tubes

The filter glass used in colour television sets contains neodymium. This helps in filtering of all colours except the three primary ones   blue green and red  rendering the screen bright and showing up a good contrast of colours.

Cerium is effective as a stabilizer to counteract browning of glass by cathode rays. Hence cerium oxide of 90% purity (CeO2/REO) is added to TV and cathode ray tube face plates. Since 1980 addition of neodymium (95 98% pure) to television tubes is practiced to cut off ultraviolet light. Consumption of cerium in the manufacture of television and cathode ray tubes is considerable. In Japan alone it is reported to be about 500t per year.

Glass for Eye protection

Glass with 2   4% CeO2 and small amount (0.5%) of praseodymium and neodymium added as didymium oxide absorbs both ultraviolet and infrared radiation. Such glass has long been used in the fabrication of glass blowers and welders  goggles. It also effectively absorbs the yellow of sodium light

Infrared Transmitting Glass

Addition of some heavy rare earths to special glass confers on it wider transmission range of infrared. At the same time such addition brings about higher chemical stability for the glass. Glasses made with mixed fluoride additives like ZrF4 LaF3 BaF2 and ZrF4 ThF4 LaF3 and germanate glass with addition of zirconium oxide and lanthanum oxide  has excellent infrared transmittance capacity.

Radiation Protection Windows

A specialized  though not a bulk use of rare earths is the fabrication of thick lead glass slabs for radiation shields and windows in nuclear installations. On exposure to a high dose of  radiation such lead glass  normally develops colour centers and gets discoloured. Addition of about 3% CeO2 prevents this type of damage and maintains transparency over long periods. Since such windows are very thick even a small amount of praseodymium and/or neodymium results in severe lowering of the transmittancy of the glass. Hence for this purpose CeO2 of 99.9+% purity is required.

Optical Glass

From the beginning of the twentieth century efforts are on to make high dispersion glass with low refraction and low dispersion glass with high refraction for optical lenses. Schott Co. of Germany developed optical glass with high barium content to become the leader in the camera world before the wars.

After the war  Eastman Kodak Co. of U.S. introduced special optical glass containing low silica and having lanthanum oxide as a component. Lanthanum incorporation in glass reduces dispersion by raising the refractive index. It also confers a high degree of chemical stability  to the otherwise weak borate glass  with respect to the action of water and acid  at the same time increasing its hardness. Prior to the introduction of lanthanum glass it was necessary for fabricators of optical instruments like cameras  microscopes  binoculars etc. to combine several concave and convex lenses made of glass of different refractive indices and dispersion power for eliminating spherical and chromatic abberation of the image. Various borate glass systems containing oxides of rare earths and of some rare elements like niobium and tantalum provide lenses with large aperture wide field of view  as well as long and variable focus.

Methods of Separation


Economically there is a limit to the amount which can prudently be spent in grinding. Technically there are degrees of liberation  which directly affect the efficiency of separation and the purity of the resultant products. The word concentration denotes the selective separation of the head feed into characteristic products. In the simple case of an ore containing galena (PbS) and calcite (CaCO3) comminution followed by concentration might produce three fractions.

Here a concentrating process has been applied which has segregated the broken ore into three products  two of which are  finished  the valuable galena and the valueless calcite and one in which the two constituents remain interlocked as particles of middlings. Processing methods are normally concerned with physical separation only  so nothing more need be done to products (1) and (3). If further treatment is decided on  product (2) can be crushed finer and given another separating treatment.

This intermediate middling  being composed partly of the sought con­centrate  galena  and partly of the waste product  calcite  should neither be accepted as a concentrate nor rejected as a tailing. If accepted  it would lower the grade and thus prejudice sales. If rejected  its galena would be lost. In all concentrating processes the decision as to further treatment of middlings includes technical and economic considerations specific to the ore.

When the ore body contains more than one economically recoverable mineral treatment is more complicated. Not only does each species contribute a middling product  but these middlings may contain more than one valuable mineral and thus lead to contamination of concentrate A by A B type mid­dlings. Treatment must sometimes be elaborated in order to reach a speci­fied degree of purity in the end product. In general terms  a multi product flow sheet provides for series concentration  as at Rammelsberg  where differential flotation is used to produce a series of high grade concentrates serially.

Particle Characteristics

If comminution of the ore is efficiently performed  each of the multitude of resulting particles acquires distinguishing characteristics which can be exploited by a suitable separating (concentrating) process. The treatment chosen in the case of the mixture of galena and calcite is applied to the pulp to make as complete a separation between the two minerals as their degree of liberation allows. It is not necessarily completed in a single step  because of the presence of incompletely liberated middlings.

These (d) products could be disposed of as a low grade concentrate and a rich tailing  re treated  or stockpiled  if better treatment facilities or higher realisation prices were expected in due course.

The characteristics of the particle undergoing treatment may combine to make the work of separation easy  or they may interfere with each other  making concentration difficult. To be usable  they must be such that at the same time and in the same pulp the two kinds of particles (concentrate and tailing) do not  to any important extent respond in the same manner to the chosen treatment.


Provision is usually made in concentrating appliances for the separation of the feed into three products  (a) concentrate  (b) middlings and (c) gangue or tailing. In gravity separation a particle sufficiently coarse to move princi­pally in accordance with its gravitational pull can be moved in one direction if clean and heavy  and another if clean and light. If it consists partly of heavy mineral  and partly of light gangue  it either remains neutral or moves feebly in the direction either of the light or the heavy particles. This depends on which of its two component minerals predominates in the locked (incom­pletely liberated) middling. If the separating appliance has three exit channels  this indeterminate middling forms a dividing band between the two finished products (concentrate and tailing) as shown in Fig. 1.

If the middling  instead of being run through the appliance in open circuit  is returned and mixed with new feed the total quantity of middling in the circuit increases until the position shown in Fig. 4 is reached.

It is thus possible to give these hesitant particles more chances to sort themselves out. If they are made to issue from the appliance in a band lying between concentrate and tailing  the separation is at the same time made more effective.

A 100% circulating load has been built up at the moment illustrated. For each eight units of new feed eight are being recirculated via the middlings band  while eight leave the system  three as finished concentrate and five as finished tailings.

While this happens  new feed comes to the separating appliance  and the middlings leaving is therefore increased beyond the 100% circulating load. This cannot continue indefinitely. A balance must be struck between the amount returned and the amount of new feed. When the operation has steadied to this state of balance  the position is again as depicted in Fig.1 save that the return circuit of middlings now provides an excellent dividing zone between concentrate and tail. As the separating appliance does not alter the degree of liberation of the middlings  something more must be

If the separating appliance is unsuitable for the treatment of the ground middling  the latter is sent to a more appropriate machine. Three important points in mill practice should now be apparent. First  it is not necessary to begin by grinding all the minerals to their fully liberated state in order to procure clean separation. Results can frequently be achieved in stages. First comes grinding  next separation into clean concentrate  clean tailing and  locked  middling (as incompletely liberated particles are called). Finally  the middling is unlocked by grinding  and retreated.

Second  the middling  or any fraction of it  can be held in a closed circuit if by doing so the work of separation is made more efficient.

Third  a true middling always needs special treatment not provided for in the appliance which has sent it out as a middling. After this special treatment it may be unsuitable for return to the sorting appliance. In the case just considered  if treatment depended on the mass of the panicle  the ground fragments should probably be sent to a machine specially adapted to deal with their smaller size.

Types of Middling

Table 1. lists a number of mineral characteristics of which advantage may be taken to achieve separation. The behaviour of a middling can sometimes be strongly influenced by the characteristics of one of its minerals while the other part of the binary system offers little or no opposition. The coarser the grinding  the cheaper it is and the greater the amount of middling made. Where binary particles are sufficiently responsive  advantage can be taken of then breaking behaviour to cheapen the working costs without losing valuable mineral. Middling particles can be associated in various ways.

A particle of type (a) would need breaking before it could be correctly graded. Since hall of its surface is gangue and half value it will respond well to surface attack such as is employed in chemical or flotation treatment Particle (c). in which the value is deposited as a shell on a core of gangue is likely to behave as a tailing in gravity work and as a clean concentrate in flotation   (d) is usually lost in flotation  but satisfactorily relieved of its value in a solvating process  (a)  (c). and (d). if the value is ferromagnetic  should respond to electromagnetic pull. Particles (b)  (c). (f). and (g) might act as gravity middlings. When the values are almost or quite masked by gangue they will he completely lost in flotation or chemical attack. Particle (g) can be a special case. Suppose the gangue to be quartz the outer enclosed mineral pyrite and the centrally enclosed mineral gold (the last being the value sought)  then the grinding treatment must expose the gold. Regrinding must therefore be far more elaborate than in usual cases  particularly if the gold is segregated in the pyrite in specks only a few microns in size. Here roasting treatment is usually preferred to further comminution.

Coarsely mineralised ores suitable for such treatment are now rare. Further  most minerals can be cheaply concentrated at a fine m.o.g. by flotation  so the need to avoid line grinding which dictated gravity concentration methods before 1920 no longer exists. If flotation is used  as it is today in most mineral dressing plants  instead of removing finished concentrates by stages it may be feasible to discard liberated gangue at each grinding stage  as in Fig. 8.

This principle may be used to give selective treatment to the middlings product at each grinding stage  as in Fig. 9.

This arrangement is a variation of the cascade principle which is used in chemical engineering to enrich or deplete an original feed by gentle stages.

The extent to which this principle of gentle upgrading is used in a given case depends on the operating difficulties which must be overcome and on the cost warranted bv extra treatment.


At this stage the student is recommended to practise separation of a heavy mineral from a light one by panning. This will familiarize him with the  packing  of sands  the transporting power of a slurry the difference in behaviour of value midding and tailing particles and the difficulties encountered in attempting to treat a long ranged feed in a single operation. A clean prospecting pan should be used. Scouring with sand usually suffices to remove old rust. The pan should not contain oil or grease. If a gold ore is to be tested  the pan should be dark in colour so as to show up the golden specks clearly. Since a treat deal of panning is done in pools and streams  the beginner should learn to squat on his heels and pan from one gold pan into another full of water  so that he can save and retreat the discarded material until he has become expert.

An excellent practice material is a  20–mesh mixture of sand and galena. A few hundred grams of this material are wetted down into water and worked into a running pulp. This is next deslimed by gentle decantation  the pan being held with its double riffle away from the operator. At no time should pulp be allowed to stream over the riffle. It should be floated out into the pool of water with a gentle swirling or rocking motion.

From time to time the pan should be tapped with the heel of the hand to aid the heavy particles to burrow down to the bottom of the fluid body of pulp. The top strata should then be panned off  using a jigging or swirling motion. As soon as heavy panicles show  the material still in the pan should be repulped and rethumped. Successive barren strata can thus be removed without loss of values. When the point is reached where most of the sand has been rejected  the decision must be made as to whether a low grade concentrate and a clean tailing is to be produced  or a high grade concentrate and a middling. In the latter case  rejected sands which now carry heavy mineral should be panned out into another holding vessel. It is not possible to make a clean concentrate and a clean tailing in one operation. The rejected tailings should be repanned to see how efficiently the work was carried out. With practice  it is possible to use panning as a rough guide in assessing efficiency of gravity treatment of sands.

The plaque  a white enamelled concave disc 11  in diameter  is used in a similar manner in the examination of fine sands.

Gravity Separation

In gravity separation the combined effect of mass and shape of the particle determines its movement relative to flowing water. In one development of this effect the water flows vertically either continuously (classification) or in oscillating motion (jigging). In a second method the ore is fed into a fairly quiet pool of dense media (water mixed with slow settling heavy minerals to form a fluid which can be maintained at a high specific gravity). Ore entering this dense medium either floats or sinks. A third type of gravity treatment uses flowing streams to effect separation. Here the pulp is carried horizontally) or down a slope  and separation depends on the rate of fall  and resistance to displacement after the particle reaches the floor of the appliance.

Methods which exploit differences of gravity require that there shall be a marked difference between the specific gravities of the value and the gangue. The material must be sufficiently coarse to move in accordance with Newton s law. Particles so small as to settle in accordance with Stokes  law are unsuitable for concentration by simple gravity methods. If centrifugal force is applied  such fine particles can sometimes be treated. Gravity methods of separation become cumbersome and inefficient when the average particle is so minute that its surface friction dominates its movement through the surrounding fluid. Exploitation of differences between the specific gravi­ties of particles below 150 to 200 mesh is difficult and usually avoided when alternative methods of treatment are available.



Crushing grinding and other words or phrases associated with the size reduction of ore and other rock are all comprehended in the word comminution. This (Truscott) is the whole operation of reducing the crude ore to the fineness necessary for mechanical separation or for metallurgical treat­ment. It is usual to make an arbitrary division of comminution into con­venient stages. Primary crushing brings run of mine ore down to a maximum size of the order 4  to 6  in average diameter  secondary crushing receives feed at  6  and reduces it to below 3/4 .  Dry crushing includes work on ore as mined  which may be somewhat moist when delivered. It is succeeded by comminution in water  arbitrarily called  grinding . Although a con­siderable amount of fine grinding is done by dry methods  this book follows usage by reserving the word  crushing  for an operation predominantly dry and  grinding  for work on a suspension of ore particles in water. One important difference between dry and wet comminution lies in the mode of seizure of the particle. In the former case the particle is large enough to be gripped between two solid steel members as they are pressed together by mechanical forces. One or both of these members moves to and from in a fixed path cycle. The rock gravitating through the rapidly expanding and contracting gap thus produced is nipped and crushed. In wet grinding the bulk of the ore is already too finely divided for a particle to be seized in this manner. It is therefore exposed while more or less free to move  to random blows. There are exceptions to this generalisation.

Machines used in dry crushing must work in dusty conditions  even when the main cause (escape of fine particles at transfer points) is dealt with. They are usually worked intermittently  to fit in with the hoisting and delivery programme of the mine. On completion of the dry crushing their product is delivered to bulk storage (in the mill s fine ore bins). From these it is de­livered at a controlled rate to the more continuous grinding and concentrating processes.

The nett energy consumed during equal reduction ratios in comminution increases with increased fineness of the ore being treated. From experimental evidence Hukki suggests an apportionment of this power consumption of 0 35 kWh/ton in primary crushing of a brittle solid  rising to 0.6 in the secondary crushing  1.6 in coarse grinding and 10 kW/h in fine grinding in the ratios stated above. These  in his view  suggest an important change in the use made of the applied power through these stages. At the primary crushing level  the results correspond statistically with the requirements of Kick s Law This phase is followed by approximate agreement with Rittinger s Law in the intermediate stage  while Bond s formulation becomes increasingly good at the finer end of comminution. These laws are discussed below. The enormous rise in power consumption when a product well below a micron (u) in size is aimed at more or less rules out the use of standard grinding techniques on the score of cost.


The main purposes are

Convenience in transport

     Production  for use without further treatment beyond screening  of graded sizes and shapes

      Liberation of specific mineral/s as a step in separate recovery from the ore

Exposure of contained values to chemical attack

Production of granular material suitable for treatment by gravity methods

Development of particles suitable for feed to froth flotation

The methods of treatment are discussed later in this book  but must be recognised from the start as depending for efficiency on correct comminution. The proposed end use should dictate the stages and methods employed. A granite or limestone ballast for railroad beds or roadmaking would have no problem of specific liberation  but would be concerned with particle shapes and sizes as these affected packing  drainage through voids  and structural strength. Preparation for the processing of an ore would call for much finer crushing and grinding  in which the cost which rises sharply when fine sands must be prepared as feed to the concentrating plant would be an important factor.

The earliest stage of rock breaking is in connection with the severance of the ore from its lode. This is performed by the use of suitable explosives applied in such a manner as to produce lumps of rock of a size convenient for handling in the mine s transport system. The interest of the alert ore dresser can well commence at this early point since the manner in which the abrupt stresses of explosion are applied to the rock in situ is one determinant of its size range  crushability  and total surface per unit volume from that point on to the next crushing stage. The lavish use of high explosive underground is to be depreciated  first on grounds of cost  second because the finely shattered ore thus produced is difficult to gather and transport  and third because the delicate reactions used in treatment are jeopardised by random exposure of small particles between slope and process control in the mill.

Ore  as broken  may range in size from lumps weighing several tons down­ward  but delivery passages  chutes  gates  trucks  and skips work best when they are not exposed to the shock loading of large pieces of dropped material  and  pack  best when the sizes transported are reasonably close ranged. Sev­erance may be followed by the use of explosive on large pieces of ore lying in the slope  by sledging  breaking down lump ore on a grating protecting an ore chute underground  or at an underground crushing station  where a jaw crusher of the Blake type is frequently installed. This machine should be set in the updraft ventilating zone so that the dust produced in crushing does not contaminate the mine air. Broken mine timber coming to the crusher with the ore must be removed. At all stages of ore treatment it is good practice to remove trump iron and other mining detritus as early as possible. The feed to the underground crusher must be adequately displayed and illuminated  to facilitate hand picking by the crusher attendant.

Big tonnages of rock are crushed for use as graded sizes of stone or homo­geneous rock for road metal  ballasting  etc. Such  dressing  has nothing to do with mineral processing  in which crushing is a stage in the liberation of values. Washing and sorting may be applied to the ore in transit and simple concentrating treatments may be used at the same time  but the main purpose in crushing is to reduce the size of the rock particles by suitable stages so that the most efficient use of force is made and unnecessary comminution is avoided. For some purposes dry crushing must be used throughout the work e.g.  in the grinding of cement  dressing of mica  talc and some other minerals. Generally comminution commences by dry crushing the ore to below a size established by tests  and finishes by wet grinding to the required liberation size or mesh of grind (m.o.g.). The changeover from dry crushing to wet grinding lies between 3/4 and 1/4  and tends toward still lower sizes with the introduction of tougher alloys  improvement in design  and better methods of lubrication. These have led to the development of crushing machinery which can withstand the severe working stresses involved when large tonnages are crushed to gravel size.

Crushing Theory

The forces used to produce fracture of a perfect crystal are of two main types. The structure is bound together by its inter atomic forces of attraction. Stress as considerable as 10  p.s.i. (pounds per square inch) is required to disrupt this bonding or  theoretical strength   which can be  calculated. If tension is applied the crystal stretches elastically until it reaches its yield point  and recovers if the stress is removed before this point is reached. Once the elastic limit is exceeded a flaw is produced  usually in the form of a minute crack  which becomes a focus for incipient fracture. In his classic paper Griffiths noted that the stress  which the crystal can thenceforward withstand  is inversely proportional to the square root of the length of the crack. This means that it will now fracture at a much diminished stress loading. During the stressing which created the imperfection  work was done to overcome the mutual bonding of the inter atomic forces. This work was stored as elastic energy  and was released as the atoms returned to their normal positions. Since a crack existed  the atoms in its vicinity were able to shed elastic energy while the crystal was being stressed. If this released energy was sufficient to overcome the weakened inter atomic bonds at the tips of the crack  it grew rapidly (at a speed of about 15 000 ft/sec). The fact that stress of the atomic bonding is focused at crack tips has been proved by polarised light studies of plastic materials.

Such a crack in solid material may start as a scratch or a surface blemish (a superficial discontinuity)  as a minute fissure in a crystallite structure  or as a defect in the atomic lattice of a crystal grain  which  by yielding  permits the start of plastic flow. Most rocks and man made materials contain such foci of weakness  so that practical strength falls far short of theoretical. Ideal glass should withstand stresses up to 10 p.s.i.  but in practice failure occurs at about one ton p.s.i.

Where dislocations exist  crystals  which would otherwise deform plastically as one plane slides freely over another  have this plasticity blocked at each such dislocation. Stress builds up at this point and the surrounding atomic bonds are ruptured. The result is a minute crack which  given a little rela­tively light further stressing  becomes a complete fracture. Once crack propagation begins it proceeds at nearly the speed of sound. A corollary is that once a crack is running at this speed  no further stress applied to the main structure can catch up with the advancing fracture.

It is thus clear that two kinds of stress can operate the reversible  which is removed if the crystal is not loaded to its elastic limit  and the irreversible in which surface discontinuities are formed or plastic deformation occurs. This last is the result of the slipping of one plane of atoms over another. The slip is local and so produces a deformed area  with one side in compression and the other in tension. Energy is thus stored ready for use in further deformation. Similar storage builds up in zones surrounding dislocations  which prevent the free sliding of one atomic plane over another. Naturally occurring rocks contain random focal points where stresses due to defects are stored till an initial crack is produced by a relatively small further stress. Further weakening factors are the explosive shock during blasting of ore and its chemical oxidation between severance and milling.

In a piece of ore at least two mineral species are inter crystallised in various patterns. The situation is far more irregular than that in the perfect and completely pure crystal with which this discussion commenced. The prac­tical application of crushing theory must therefore add statistical methods of testing and an empirical approach to the considerations outlined above. No two lumps of ore are precisely similar. It is common to think in terms of such varieties of disruptive force as compression  tension  shear  torsion  abrasion and shatter. The last four are compound forms of the first two. When a beam suspended at its ends is centrally loaded till it bends  the lower part is in tension while the upper part is in compression. When seizure of a piece of rock occurs in such manner that the seizing forces move in opposing directions while the rock is prevented from rolling  tensile shear predominates  with local compression where its high points are gripped. Most crushing force is applied compressively. A piece of rock can be thought of as either a column or a beam loaded beyond bursting point. A piece of ore sufficiently large to be gripped forms a short pillar between approaching faces and is loaded to failure. If the ore bridges the crusher faces  beam loading results. With the exception of the dynamic class of crushing device  in which the par­ticle becomes a projectile in a fast moving gaseous stream and acquires sufficient velocity to cause it to shatter itself on an impact plate (special cases which will not receive further attention in this book)  application of stress by the crushing device is very slow in relation to the rate of propagation of a running crack. This is true even for such impact crushers as stamps and hammer mills.

Many attempts have been made to establish crushing principles on an un­assailable basis of fundamental law. Kick s  law  which Gaudin rightly says should have been regarded as a postulate states that  the energy required for producing analogous changes of configuration in geometrically similar bodies of equal technological state varies as the volumes or weights of these bodies  (Stadler). The Rittinger law  which should be regarded also as a hypothesis  states that the energy necessary for reduction of particle size is directly proportional to the increase of surface. If the sole force operating to produce crushing was used to disrupt molecular bonds along planes  and thus to produce completely severed new particles  Rittinger s law would be in line with current physical concepts  but this is an over simplification.

Something of the difficulty facing the research worker in the field of commi­nution may be appreciated when it is recalled that four types of binding force are recognised in fundamental physics.  Taking nuclear force (that binding a proton to a neutron) as unity  then electrostatic force (proton to electron) is 10  nucleon decay force (emission of  particle) 10 and gravitational attraction indefinitely weak. Nothing is at present known of the balance between these forces  which must be upset for cleavage to occur. Crystal study shows  for a homogeneous material  three types of imperfection  which affect resistance to shear. Micro defects are lattice imperfections due to irregular ion distribution. They probably have a sub threshold effect on comminution  though as will be seen when the physical chemistry of mineral surfaces is studied  they modify flotative reaction. Macro defects are in­cipient strain areas  flows or discontinuities in an otherwise regularly repeated lattice structure  which in the perfect crystal would be an orderly multiplication of its unit cell.  Mosaic defects  are typical in crystals in which orderly blocks (type 2) are constituents of the overall imperfect particle. Fortunately for the mineral engineer  these considerations are of less importance than the empirical approach.

Basic research on the crushing characteristics of multi phased rock such as a piece of ore is hampered by the fact that no two pieces are alike. Repetition tests cannot therefore offer reliable evidence. Since comminution is the most expensive part of treatment and also exerts a critical influence on both concentration procedure and percentage recovery of values  empirical test procedures have been developed for use in each specific problem and for each specific ore body. The main practical interest in a given case centres on two factors. The first is the manner in which the association of different mineral particles in a piece of rock can best be disrupted. The second is the effect on subsequent treatment of any new development in grinding methods.

A new approach to the theoretical consideration of crushing has been made by F. C. Bond. He commences by challenging three assumptions which weaken the classical theories. These neglect the work previously done on feed particles under examination  although such effort is part of the total input of crushing work. Secondly  the old theories were arrived at from study of breakage of cubes  not of the irregularly shaped particles handled in milling. Thirdly  these theories equate useful work input against energy increase on breakage and neglect the energy released as heat as  wasted  or external to the problem. Bond proposes a theory intended to give con­sistent results over all size reduction ranges for all materials and machines.

 It appears that neither the Rittinger theory  which is concerned only with surface  nor the Kick theory  which is concerned only with volume  can be completely correct. Crushing and grinding are concerned both with surface and volume  the absorption of evenly applied stresses is proportional to the volume concerned  but breakage starts with a crack tip. Usually on the surface  and the concentration of stresses on the surface motivates the formation of the crack tips.



The Duty of the Section

In primary crushing the largest lumps of ore mined must be dealt with. In secondary crushing the maximum sized piece is unlikely to exceed 6  in average diameter and some of the unwanted material coming from underground has probably been removed. The feed is therefore easier to handle. The crushing  machines need not have so wide a gap nor so sturdy a construction. The transporting arrangements can be less robust  since the large pieces of rock have now been reduced to more manageable fragments.

Washing and sorting  if practised  may be done in the primary section  but is more usually combined with secondary crushing  where the rock is smaller and more easy to handle. Secondary crushers are usually arranged in series with the primaries  so they must be able to handle similar loads. Their main task is to reduce the ore to a size suitable for wet grinding. It then goes to the fine ore bins  which must have sufficient storage capacity to receive all the ore accepted for treatment and to keep the plant running continuously  although the mine only delivers its ore periodically.

Layout and Equipment

A generalised secondary crushing scheme in which minus 6  rock is sorted and reduced to minus 1  in one operation or  pass  is shown in Fig. 1. The numbers in this flow sheet refer to

     Transport and feed regulation from primary crushers to screen 2. The feeder stops  starts or modifies rate of delivery and the guard magnet removes magnetic iron ahead of 4. It could be used at 3 instead of I  or be dispensed with if hand picking were used on 3 to remove iron. This might be done where the danger of damage to 4 lay in the passage of manganese steel or other non magnetic and uncrushable material.

Separation of finished undersize from plus 1  rock which is to be crushed. A robust screen system is used.

     Sorting  picking or transporting belt (belt conveyor). This delivers to 4  and may elevate the ore in transit so that the crushed rock leaving 4 falls by gravity to I. The   6  + 1  ore is of a convenient size for handling  so removal of waste and detritus is possible. The side arrow shows that provision for collection and disposal of finished waste is then needed.

Secondary crusher  set to 7/8  By making the set a little below the I  screen aperture re circulation of near sized waste is kept down

     Conveyor belt or chute returning crushed ore to 2.

     Conveyor belt  perhaps equipped with weight recording equipment (weightometer) and automatic sampler  which delivers screen under size to storage.

     Fine ore bin/s  where ore accepted for treatment is received and fed at a controlled rate to the plant.

         The purpose of the weightometer (6) and sampler is to record the tonnage accepted for treatment and to sample it for assay grade  moisture and particle size. Process control requires knowledge of dry tonnage treated and the values contained in that tonnage.

If sorting or dense media separation is also practised  it will probably be introduced in this section. In the case of material requiring fine crushing by dry methods  special equipment of kinds not considered in this book may follow. The fixed path crushing machines discussed in this chapter normally deliver to the bins ore crushed to below between 1  and ½ . Closer settings are possible  particularly if the secondary crushing is performed in two stages so as to reduce mechanical strain by keeping each machine s reduction ratio below 7 to 1. The considerations which influence lay out and installed capacity include the crushability of the rock itself  the question of waste elimination and the maximum rate of delivery of ore for immediate handling. During the mine s transport period the fine ore bins are filled  though throughout the mill s working day (usually 24 hour) they deliver steadily to the next stage of treatment.

Secondary crushing is today characteristically performed  dry . This word is used relatively  as the ore may be moist  or may be wetted during washing operations or when water is run in through a crusher to prevent the build up of clay. (This last is bad practice  and may lead to mechanical trouble.) The machines mainly favoured are modified forms of gyratory crusher  though other appliances  such as crushing rolls  are in limited use. Beater mills  with their variants (hammer  whizzer  pin disintegrating and ring roll types) deal with large tonnages of coal  asbestos  limestone and easily broken mineral. The gravity stamp  though worked with added water  is briefly described in this chapter  since it is a fixed path machine. It is obsolescent  its place being taken by the rod mill described later  but is still in use in a few older plants.

The Symons Cone Crusher

As with the gyratory  crushing results from interaction between three essential parts. The important difference is that in this case the spindle (1) is not hung from its upper end but is supported in a universal bearing below the gyrating head or  cone  (2). Normals to the arc through the universal bearing carrying the breaking head intersect at O. This break­ing head gyrates inside an inverted truncated cone (3)  called the bowl  which flares outward and thus allows for the  swell  of the broken ore by providing an increasing space for it to enter after each  nip  has released crushed ore for a further drop. Two types are made  the standard and the short head. These differ chiefly in the shape of the crushing cavities. Standard crushers deliver a crushed product varying from ¼  up to 2½  usually in open circuit and can be fitted with fine  medium  coarse  or extra coarse crushing cavities. Short head crushers have a steeper head angle  a longer parallel section between cone and bowl  and a narrower feed opening. They deliver a crushed product ranging from 1/8  up to ¾   and usually work in closed circuit as in Fig.1.

The springs then return the bowl to its correct clearance. While this happens or when choked with clay  the Symons crusher is apt to let oversize escape. Such clay is sometimes dealt with by introducing water with the feed. Better practice is to remove it by washing. It is dangerous with any dry crushing machine to risk the entry of abrasives into its bearings or bevel gears.

The springs which hold down the bowl yield when the load is too severe. It is therefore usual to run the cone crusher in closed circuit with a screen  thus ensuring that  tramp oversize  is returned for further treatment. This recirculated material  which may include fragments of non magnetic steel not removed by the guard magnet  may call for removal by special methods.

With some ores there is a tendency for extra tough particles to  spring  the crusher at a slight oversize to the set. (In passing  it should be noted that in all closed circuits there is selective retention of one fraction of the ore stream which may call for special measures.) A simple solution

Ratio of reduction is controlled by screwing the bowl up or down by means of its capstan and chain.

A 2 ft. standard crusher receiving  2 3/4  rock is rated to deliver 15 tons/hour ¼  open circuit  or 60 tons/hour when reducing  4  feed to ½ . A 2 ft. short head crusher receiving 13/8  feed and delivering 1/8  in closed circuit has a capacity of 6 tons/hour. When reducing  2  feed to  ½  product the capacity is 20 tons/hour. The coarser 7 ft. standard crusher has a capacity of 900 tons/hour when reducing  18  rock to 1½ . Such general figures  taken from manufacturer s literature  would require checking for performance on samples of the specific ore if it was proposed to work a crushing plant at full capacity.

Gearless Gyratories

Several crushers are marketed which avoid the complication of bevel gearing below the rock treating section  by making the drive a direct extension of the rotor of the electric motor. The Newhouse  which is hung from cables to absorb the vibration of its running  has the motor above the grit zone. It is run at 500 to 600 r.p.m. with an eccentric throw of about ¼ . An advantage claimed for this design is that  by avoiding the use of gears  little power is consumed when the machine is  idling


Although much work once done by rolls has been taken over by cone crushers  these machines still handle a considerable tonnage. Standard

The method of feeding is important. Unless the entering ore is spread evenly over the whole width of the rolls  partial wear occurs  causing the surfaces to become grooved or flanged. The Taylor heavy duty crushing rolls incorporate a  fleeting  mechanism which causes one cylinder to move to and fro on its axis  thus reducing this type of wear. A good practical rule is to arrange the feed so that some ore falls outside the crushing area at each end. This helps to even wear over the full width of each roll. Another is to raise the feeding device so that ore arrives on the rotating surfaces at their peripheral speed. This gives the best conditions for seizure. Rolls can only work as  arrested  crushers if lightly fed  because the breaking ore swells in volume as voids are produced  at the same time as the particles fall into a more restricted space. If not  starvation  fed  rolls are choke crushers  ore grinding on ore. Unless rolls of very large diameter are used  the angle of nip limits reduction ratio  so that a flow line may require coarse crushing rolls to be followed by fine crushing rolls. Although the floating roll is only supposed to yield to an uncrushable body  the choked packing of ore in the crushing throat sets up so much pressure that the springs are usually   on the work  during crushing  and a moderate proportion of unfinished material is jet through. For this reason  rolls should be worked in closed circuit with screens wherever control of maximum particle size leaving the crusher is important.

Hammer Mills

In these machines  which can be used either as primary or secondary crushers  the breaking force is mainly due to a sharp blow applied at high speed to free falling rock. The moving parts are beaters (hammers  rec­tangular plates  hanging bars or heavy metal rings). They move in a more or less  fixed circle of swift rotation  though they are loosely suspended from pins on discs mounted on a driving shaft  inside a robust stationary casing with a grid through which broken undersize leaves the mill. The beaters weigh from a few up to 250 lb.  and the larger machines can work on feed as coarse as 8  cube. Fracture is chiefly produced by the flailing action as the beaters hit the ore as they spin at from 500 to 3 000 r.p.m.  though part of

Both uni directional and reversible hammer mills are manufactured. The former can be used over a wide size range of soft or friable material  in both primary and secondary crushing and the latter in secondary crushing. Rever­sal of the rotor direction obviates the need for turning the hammers round. A hammer mill with its breaker grid arranged like a miniature belt conveyor has been developed for sticky or wet feed liable to clog a fixed grating. Material which fails to fall through climbs against the down running new feed.

Another type of hammer mill  the  Impactor  is designed to obviate stoppage for hammer adjustment. The rotor is reversible  and end wear can also be taken up while running  by movement of the anvil blocks which regulate the set of the mill. Elimination of a retaining grid makes the machine able to cope with frozen or sticky feed and aids quick passage. The velocity with which the blow is struck in impact crushing is the main deter­minant of the severity of shock loading. The vertical distance of free fall as the feed enters is therefore a control factor. Care in arranging a suitable dropping height influences product shape  size and production of fines.

Among the minerals broken by this type of milling are limestone  spars  gypsum  shale  clays  coal  asbestos  gravels and rock required for ballasting or concrete aggregates. The products are characteristically sharply frac­tured. Moisture affects efficiency adversely  and wear on the beaters is heavy with abrasive material. Product size is controlled to some extent by varying the escape grid apertures  clearance and speed. Unlike the dry crushers hitherto considered  these machines are impact breakers.


Optimum Grind

Optimum size of release from the grinding circuit into the concentrating section is determined by technical and economic considerations. The finer an ore is ground the more this grinding costs. Up to a point  finer grinding usually results in higher recovery of values  but beyond this over grinding leads to poorer recovery. Optimum grind defines  the mesh of grind at which a maximum profit is made on sales  when both the working costs and the effect of grinding on the recovery of values have been brought into consideration. Such an optimum point is determined in the first place by test work in the laboratory when the flow sheet for a specific ore is being prepared. It then becomes the operator s aim to achieve maximum through­put at this optimum grind. With care and attention it is frequently possible to improve on the figures obtained in preliminary testing.

 Optimum grind  or  release mesh  refers to the sizing analysis of the ore particles finally leaving the grinding section. In its simplest form it can be specified as  say  a 100 mesh grind  meaning that substantially all the particles in a carefully taken sample pass through a 100 mesh screen. A more exact optimum might call for  say  minus 5% on 100 mesh and plus 85% minus 200 mesh  indicating that a little coarse and rather light gangue stuff  may safely be allowed to leave the circuit  and that the needs of the concen­trating section of the plant will be best met if overgrinding is avoided by stopping the process when 85% of the ore  by weight  passes a 200 mesh screen. Such a specification calls for care in order to bring all the material  undersize as well as oversize  between certain size limits so far as skillful control permits. It results from the scientific application of forces now to be described. Careful grinding preparation develops the latent characteristics of each mineral species it unlocks from the ore in the form of individual particles. It thus simplifies the work in the concentrating section and is of great practical importance.

When ore characteristics vary from section to section of the mine it may be found economically desirable to avoid mixing  so that each type can receive specialised treatment. Separate bins are provided in such a case. Ores may also be bedded so as to be drawn in a uniform grade through one grinding circuit. In a small operation the dispatch from the mine of markedly different types of ore should be so regulated as to provide a steady run of mill head feed.

Applied Power

Of the electrical power fed in  a loss of the order of 10% occurs in the motor  and between 10% and 15% in the gears and mechanical friction of the mill. The balance is available as  useful power  as kinetic energy in the tumbling crop load but the fact that it is available does not by itself lead to its efficient use. If the crop load is not properly constituted  part or all of this kinetic energy will be wasted by conversion to avoidable heat  sound  and ground up metal.

Fig. 1. shows torque (in the sense of useful power) to be nil at zero and also at critical speed. At critical speed (plus the extra speed necessary to compen­sate slipping of the crop load and hold it by centrifugal force) grinding stops.

Such a position is largely theoretical  since it could not be reached in a normally charged tumbling mill. The relationship between centrifugal force and its radial tangential thrust at various depths along the mill s radius passing through the crop load  precludes such seizure. The many inter­acting factors at work in the churning charge tend to confuse the picture of its dynamics. To clarify discussion  consider the case (admittedly over simplified) of a variable speed mill with smooth liners  no ore or water  and a load of steel balls of one size. As it starts very slowly from rest the load surface tilts until the slope is reached where load stability fails and the top layers of balls slide down. Neglecting slight irregularities in breaking away  the power draft is steady for a given speed when the stable displacement is at its maximum. This draft (ignoring mechanical loss outside the mill shell) corresponds with the displacement of the centre of the turning load from the vertical diameter by a distance a along the theoretical path (zero to critical) in Fig. 1. As the mill speed is increased this dis­placement reaches its maximum. With further speed increase the distance a begins to recede and power input falls. The concept of freeze up at critical speed is not valid  in view of three main forces at work  of which centrifugal fixation is only one. First  before this stage is reached any peripheral balls rising clear of the down slipping load after it has passed the plane of its horizontal diameter fail to maintain tangential direction because they are now acted on by gravity. Losing contact with the shell they take a falling trajectory to the down running side of the mill. In the course of this they collide either with other balls  loosening the upper part of the charge  or with the shell itself  thus transferring part of their kinetic energy back to the shell from which they had received it. This acts against the input of new energy once a balancing peak of flight has been passed. Second  the packing structure of the charge changes steadily as mill speed increases. At rest  ball rested on ball and voids between these spheres was at its minimum. With rising speed the core of the charge  and also the upper layers  are loosened so that the volume increases. For a given mill speed there is a critical volume of charge at which the centre of mass a is at its maximum displacement from the vertical diameter. If the volume of the charge had been sub critical at this speed it would have been possible to increase the out of balance loading by adding more balls  and the power draft would have risen. Similarly  it would have been possible to increase speed without increasing charge to obtain the same effect. If  on the other hand more balls were added  or speed were increased to the super critical point  the combined effect of the reduced out of balance dead loading and kinetic impact of falling balls on the down side of the shell and the toe of the load would be to reduce the useful input of kinetic energy to the system  and the current or wattage drawn by the driving motor of the mill would fall. The third force  the consoli­dation of the charge by centrifugal force  is modified by the first two.

The practical crop load is  of course  a mixture of grinding media of various sizes and shapes  perhaps even of varied density  since steel and large pieces of ore may form part of the crushing bodies. Next there is the ore  partly a new feed from dry crushing and partly a return load of partly finished sands from the closed circuit. Further  there is the make up water added at mill head  which modifies the plasticity of the tumbling load in accordance with its specific surface and the percentage of water used. Finally  there are the liners  of various types  shapes and degrees of wear  with grip modified by changes in the amount of slime or slippery sulphide mineral anointing their surfaces. Again  over simplifying somewhat  the crop load can be pictured as a loosely plastic body being continuously moulded into shape by tumbling action and influenced in its mass cohesiveness by the frictional hysteresis of its components. In most operating circuits the mill speed is fixed  but in all the crop load varies slightly in volume with the grindability of the ore. Since only  finished grind  pulp is allowed to leave the grinding section increased resistance of newly entering ore results in increased retention in the closed circuit  part of which takes place in the crop load. If the load then goes super critical (in terms of the above discussion) less grinding power is available  less grinding is done and the overcharge increases. If more easily ground ore comes to the mill the charge is diminished and too much energy seeks too little ore in the sub critical loading which follows.

An equation for  best operating speed  (n) in terms of internal stability of crop load and frictional grip from the shell has been proposed by Davis.

In a ball mill the rate at which power is converted into kinetic energy is fairly steady  but in a rod mill it varies somewhat abruptly owing to entangle­ments  hold ups  and momentary seizures of the rods as they turn.

If the kinetic energy is correctly applied  a maximum of properly finished ore results  but whether correctly applied or not  the power continues to enter the mill so long as the crop load is being dynamically held out of balance. Unlike nearly all other comminuting appliances  the grinding mill uses about the same amount of power all the time it is running. Substantially all the kinetic energy is finally dissipated as heat  which warms the transient pulp. In the endeavour to find a fundamental expression for grinding efficiency it has been suggested that  since new surface is proportional to grinding energy (Rittinger s law)  and since this involves the creation of new surface tension or surface energy on the newly sheared particle surface  the efficiency of grind­ing is measurable as the proportion of useful power to new surface. Calcu­lations along these lines have produced several sets of figures  all agreeing in the conclusion that grinding efficiency in the ball mill is very low well under 0 3%. For practical use  industry needs something more concrete to measure the efficiency of its daily operation. Performance in the plant can  most conveniently be judged by its relation to some selected standard of throughput. There is no good direct way of measuring the surface energy of solids. In any case the increase in surface energy of the particle which is due to the transfer of some input (grinding) energy to the newly created area is probably only part of the total rise in energy. No method exists for assessing the internal changes in energy level (physical  chemical  and elec­trical) which accompany comminution. Efficient or not  the tumbling mill is the best machine at present available for the work of grinding  which is the most expensive cost per ton item in the flow sheet.

Rittinger s law has been the subject of much research  and may be regarded as a good approximation. The operator will work on sound lines if he thinks of grinding force as resulting in new surface (in an efficient operation). The next step is for him to ensure that

As much of this new surface as possible is created on particles neither too big nor too extravagantly small for treatment in the concentrating section

As few as possible of those particles shall consist of steel abraded from the balls and liners 

      As many as possible shall consist of value containing ore.

This mental approach enables him to see the problem of grinding realistically  and to translate his vision into effective operational control.

 Useful  or Net Power

The kinetic energy generated in the crop load by transfer of driving power through the mill liners should be maintained at its peak value  in order to obtain the maximum amount of grinding from the system. This maximum draft is achieved by means of a correct balance between four main factors

Speed of mill rotation  expressed as % critical speed.

Liner grip  notably of body (horizontal) liners.

Constitution of crop load or charge C (media  ore and water).

Volume of C under running conditions.

First consider item d  a and b being fixed and c varying only as to ratio of grinding media (m) to feed (consisting of new ore  classifier returns and mill head water). Take as the starting point  peak power draft  with dis­placement of C to the rising side at its maximum unbalance. If C now in­creases part of this increased volume is re balancing the load by overspill to the down running side. It thus reduces the unbalance of C and also feeds more driving energy to the shell liners. This transferred energy does not  of course  increase the total input. The reduced unbalance does  in fact  result in a power drop  registered on the ammeter of the driving motor or motors. However deployed  the net input power is proportional to an exponential value of new surface produced  and any reduction of this input leads to a corresponding diminution of useful grinding. This is why the mill must be run with a maximum input of useful power in order to maintain peak effi­ciency.

Some qualification is desirable at this point. The purpose  of grinding is not solely technical. It must contribute to the maximum overall profit  which depends on a balance between all processing costs and the grade and percentage recovery of concentrate in the best condition for further use. The size analysis of the solids in the mill discharge has an important bearing on subsequent treatment  and grinding is a major cost item. This frequently raises the operating question of shattering versus abrasion in the grinding section. The fierceness of shatter at constant speed will obviously be reduced as C increases in weight and volume beyond peak displacement of its centre of mass. Slippage will then increase  balls be more blanketed and impeeded in falling  and the toe of the charge will be more abraded and less hammered. The question always arises   What kind of grind is best for a specific ore  treated by a specific method?  The reader should re read this section after the chapters which deal with various methods of concentration have been studied.

A change in loading volume due to variation in retention of ore upsets the balance between m and C  and also alters the frictional characteristics of C. If C increases  shatter is reduced and the coarser particles of ore are less adequately reduced to sizes which m can seize and abrade. At the same time the specific surface of C decrease    though no change  has been made in the carefully controlled solid liquid ratio which K being maintained in relation to optimum specific surface. From this point there is cumulative deterioration in the efficiency of comminution. Slowly the mill discharge size analysis increases its percentage of coarse material. This discharge is returned by the closed circuit classifying system in increased volume though there has been no increase in new feed of ore. Thus  both C s texture and volume are changing from their optimum balance at an increasing rate. This readily observed connexion between reduced power consumption and loss of grinding efficiency (which may have adverse effects right down the flow line if it leads to overloading with wrongly ground material) underlines the vital importance of a well controlled grinding section.

Hems a and b may now be related to this discussion. When mill speed is increased without any other alteration C is reduced  and vice versa. This fact was brought out clearly in the Vassbo experimental work referred to earlier. There  the mill s best performance in overall terms was found to be 60% critical speed. Taking this 60% as the index of 100% efficiency  each increase of 1 % of speed was accompanied by a drop of 1 % or so in efficiency. At 80% critical it was necessary to use 25% more power to do the same amount of grinding. This extra power was obtained by increased unbalance of the crop load. The point is significant  since it shows that maximum unbalance (accompanied by maximum draft of useful power) cannot be taken in isolation as the criterion of efficiency. In this case it showed that the mill was too big for its job and in consequence its diameter has since been reduced. Maximum power draft must be related to the required finished grind  and achieved by the correct composition and volume of C. In the Vassbo operation autogenous grinding and a variable speed mill were used  with automatic linkage between speed and change in the volume of returned circulating load. With ore as m change in size composition of m components is far faster than with steel balls  and variation is easier to arrange and ob­serve. The work was done on a full working scale and tied in with the sub­sequent treatment  so that effects on recovery could also be seen. The inter acting factors thus revealed should affect grinding research and develop­ment.

A few further points may now be noted as accessory to the above discussion. When the volume is steady  the net power is highest with the interstices between balls full of ore and lowest with them full of water  which has a much lower density and therefore reduces the total crop weight. Power used is higher with plenty of  sharp  sand in the crop than with slimy sand only  since the extra friction helps the liners to grip the load more firmly and raise it higher. It is higher with a low discharge mill  because pulp rises centrifugally on the rising side but can escape near the periphery of the grate  whereas in the high discharge mill it can only overflow from the trunnion  so that a larger volume of pulp must be retained on the falling side of the mill. A drop in the ammeter reading of power input to a low discharge mill might show that the discharge grates were partly clogged.




Though most ores are reduced by wet grinding before being processed  some can better be ground and treated dry. Many minerals and synthetic substances require size reduction only. Other grinding problems arise in which chemical instability  contamination  corrosion or risk of explosion call for special precautions  such as milling in an inert atmosphere or one where moisture is undesirable or must be removed. In an arid country the chronic shortage of water may dictate the use of dry grinding methods. Where a dry end product is called for and can be processed up to the re­quired state without the use of water  dry grinding is to be preferred. Among the raw materials thus treated are asbestos rock and  crudy   coal for pow­dered fuel  cement clinker  talc  metal powders  drugs  and chemical salts. In addition to open and closed circuit grinding  batch treatment is frequently used. In this method  grinding media and feed are loaded into the grinding mill and worked dry until the desired state of attrition has been achieved. The product is then discharged.

In the treatment of ores by chemical methods  such as the cyanide process  experimental dry grinding has shown promise. When comminution is followed by froth flotation it is usual to protect the newly developed mineral surfaces  and this is best done by grinding under water to which any required protecting chemicals can be added. The technical applications of dry grinding in mineral dressing are at present limited by this consideration.

Fixed path Mills

Taggart classifies dry mills into two groups   in which the comminuting elements are relatively few and follow definite paths (fixed path mills)   and those in which  the elements are multifarious  and not constrained as to individual paths (tumbling mills) . The latter do the bulk of industrial dry grinding  but the former  of which there are several types  handle an im­portant tonnage.

Burr mills range from the old fashioned grindstone  originally used in grinding cereals  to vertical types. Two discs of stone  either horizontal or upright  are rotated in opposite directions  or worked with the lower one fixed and the upper revolving. Feed is central and finds its way along grooves in the stone faces  maintained by stone dressing  to a peripheral discharge. The material is ground by attrition during its journey  being dragged between the stone faces. Soft rocks such as clays  barytes  talc  lime and limestone and gypsum  are treated in these mills. Feed is minus ¼  and discharge can be as fine as minus 200 mesh. The mills are used for grinding material not likely to be injured by frictional heat  and also where staining by iron must be avoided. Care must be used to keep hard or uncrushable material out of the feed. Developments of this principle include vertical disc mills with steel grinding faces. The laboratory disc grinder is widely used. Its discs can be parallel or slightly offset to one another  the latter arrangement reduc­ing choke and improving throughput. Capacity is low.

Hammer mills and rolls were discussed with intermediate crushers. They can also be used for fine grinding. A special application of the hammer mill is in the crushing of asbestos. The requirements are unusual  in that the material  as mined  carries the valuable fibre sandwiched between layers of shale. Hammer mills with heavy manganese steel plates are used to  fiberize  the blocky fibrous rock. The beating action opens the fibres and loosens attached shale. This is thus reduced to a fine grit which can be screened away.

The Raymond Impax Pulveriser introduces material through a roll feeder to the grinding chamber  which is swept through by an air stream. The air entrains finished material and dust and carries them to a collecting cyclone and dust collecting chamber. A further refinement in fine grinding by hammer milling is reported by Robertson. This is a two stage mill  in which the runners in the second chamber have a higher peripheral speed at which they complete the work transferred from the first chamber.

The jet pulveriser or microniser carries a feed of  1/8  material in air or steam at a pressure of about 100lb./sq. in. This streams out through suit­able circular expanding chambers from its tangential delivery. Extremely fine grinding results  partly by mutual jostling between the solid particles  partly by contact with the chamber walls and by pressure release. Finished fine product is discharged from the centre at sizes varying down to one or two microns. The capacity is good  but wear is heavy and the mill is limited to specialised work.

Edge runners resemble Chilean mills with the mode of motion reversed  since the rollers remain stationary while the disc on which they bear revolves. The rollers are spring loaded. They  and the disc  may be made of ceramic material. Plain iron and perforated iron are also used. These mills are used to treat clays and ceramics. One form is the German Loesche Mill  from which the Hardinge disc roll mill has been adapted. In this two conic section rolls ride above a revolving horizontal disc. This disc table  on which the ore arrives centrally  runs at a speed just below that at which peripheral discharge of crushed material begins. A dam forms round the circumference over which the discharging material is pushed by arriving feed  and falls into a classifying air stream which lifts finished mineral and returns anything coarser to the grinding disc. In the Lopulco mill there are either two or three spring loaded conic rollers. These mills range in output from 1 to 50 tons/hour and are designed for low  medium or high speed. There is external provision for adjustment of clearance between the rotating table and the rolls  which cannot make direct contact. The feed receives both loaded crushing and shearing attrition. An exhaust fan maintains air sweeping and removes finished product. In addition to its wide use in pro­ducing powdered coal this mill grinds a variety of softish minerals substan­tially through 100 mesh. The list includes gypsum  lime  phosphate rock and various industrial earths.

The older pendulum  or roller mills include the Huntington and the Griffin. In both  one or more pendulums revolve inside a wearing ring against which they bear owing to the centrifugal force set up by their rotation. Material trapped between roller and ring is ground till it escapes through guarding screens  set peripherally. In the Williams mill three to five rollers are pressed outward in similar manner  but the mill is swept through by a current of air which carries to a collecting cyclone or air filtering arrangement all particles small enough to be borne along. The air or gas can be preheated in a furnace  and natural draught is aided by an exhaust fan above the grinding compart­ment. In the Raymond bowl mill  which is used for producing pulverised coal  the bowl rotates against spring loaded mullers and finished material is removed by a current of air. The Babcock and Wilcox machine has balls rotating in a horizontal grinding ring  where they press on and pulverise material fed down into the ring  the finished product dropping by gravity to an external classifier. In a variation  the mill is swept by air or gas  hot if necessary  discharge being upward. Where air sweeping is used  hot air can can be used for the purpose of drying the feed.

The Vibrating Mill

Intermediate between the fixed path and the fully tumbling mill is the vibrating ball mill. This has not changed importantly since its prototypes were developed in the United States and Germany in the pre war period. An industrial model  now in increasing use  as described by Paricio is nearly 6  high  9½  long and 7  wide. It weighs six tons and is vibrated by the un­balanced rotation of eccentrics driven by two 50 h.p. motors at 1 200 r.p.m. The layout consists of a grinding cylinder rigidly attached to eccentric mechanisms in independent horizontal cylinders parallel with it. This assembly is mounted on four sturdy vertical springs and vibrates with an amplitude of 3/4 . In action the grinding media have a vibrating period of about 1 160 r.p.m. and occupy about 80% of the mill s volume. The load tends to rotate. For suitable ores this system has proved versatile and cheap in standing space  cost of installation  weight and maintenance. A 30  dia­meter mill reduces minus ¼  soft and friable feed to 99% minim 325 mesh at a rate of 2½ tons/hour. Feed and make up media enter through a dust tight fexible spout and are discharged via a retaining grate. This mill can also be worked wet  and the retention time can vary from one minute up. When used for batch grinding two receiving tanks are used and the material is worked through from one to the other and then back until the required fineness is reached.

Vibration milling has met the need for grinding metallic powders to sub micronic sizes in an inert atmosphere  and handles such inflammable elements in this way as aluminium and magnesium. Capacities quoted range  on limestone and using steel balls  from I ton hourly in the 15  diameter mill to I5 tons in the 42 . an 80% mimus 4 mesh feed being reduced to 80% minus 200 mesh. Where iron in the product is inadmissible  alumina balls and special linings are available. Materials handled commercially include tungsten carbide  aluminous nickel  silicon carbide and iron oxide.

Tumbling Mills

Dry ball mills have much in common with those described in Chapter 5  but there are marked differences in the method of operation. Continuous and batch grinding are practised  sometimes with rods as grinding media  but more usually with balls or autogenous crushing bodies.

Since batch treatment mills are not limited in design to types which can be fed at one end and discharged from the other  a variety of shapes are in use cylindrical  conico cylindrical  oval  polygonal  and even cubic. The liners and crushing media can be made of iron or ceramics  according to the requirement of the work. Feed and discharge are usually made via an aperture in the shell  a screen being placed over it to retain crushing media during discharge. The mill is charged  run for a suitable period  stopped  opened  and emptied.

The ball load is kept lower in dry than in wet grinding below 40% to avoid  over carry  at cataracting speeds  which would cause the flying balls to hit the down running side.

One of the older methods of ensuring fineness in dry grinding is to divide the ball mill into two or more compartments. These are separated by grates which retain the material in the compartment for which crop loading is most suitable until it can pass through the grate and be elevated by scoops to the next. The Hardinge Tricone Compartment mill uses the back slope of it first (conic) section to promote media segregation and makes it possible to dispense with the third section shown in Fig 5.

In an air swept single compartment mill  which has gained considerable favour in the preparation of pulverised coal  the difficulty of maintaining a gradient through the long cone of the Hardinge is partly overcome by using a retaining grate toward the discharge end. Behind this  lifters pick up the undersize and drop it into the classifying stream of air sweeping through the mill.

Pilot tests show one step comminution to use less power than a. conven­tional crushing and grinding sequence and to favour medium and fine­ grained ores which require finer finished grinding  where air sweeping is more effective. Ore tends to disintegrate selectively along grain boundaries and to produce mineral species at their natural grain size. Any such effect assists subsequent concentration. Liberation appears better than with rod milling and wear is less. If selective grinding accompanies quick passage through the mill the later treatment benefits by having less over ground particles to deal with. The Swedish tests suggest that when working for a coarser liberation mesh  air sweeping is not fully effective. Tests at Doorn fontein showed higher recovery both of gold and uranium after 24 hour laboratory leaching of minus 200 mesh banket ore crushed direct from run of mine feed than of various samples drawn from the conventional flow line. With a steady increase in the chemical extraction of values from their ores this improvement  which again suggests disintegration along grain bound­aries  attracts the interest of process research.

Under normal weather conditions  Weston considers a moisture content in the feed tolerable up to 3½%  but in sub zero temperatures external heat to be needed from 1½%. Installed capital costs are lower than for com­parable wet grinding and maintenance is much lower.

It is used in dry grinding in the same general way.


The moisture content of the feed is an important factor in dry milling. If nothing is done to classify a circulating load  a very small percentage of water (½% to1¼%) ruins fine grinding. If the circuit is closed through fine screens  a little more moisture can be tolerated without clogging the screen apertures  unless the material is soft and clinging. The temperature of the circulating air rises as the ore is milled  and its humidity increases when it picks up moisture from the passing feed. Enough air must be bled off to avoid saturation and thus ensure that this transfer of moisture can continue. The heat requirement for drying changes with the season  and must be supplied by warming the air when necessary. The water content affects grinding  cling in the crop load  cushioning  cataracting  coating of crushing media  and mobility and classification of partly finished material in the closed circuit. As with wet grinding work  the mill capacity required for a given rate of throughput depends on the grindability of the ore and on the sizing analyses of feed and finished product. The finer the grinding is taken  the lower will be the tonnage treated. Removal of finished material is very important to high capacity. If it is allowed to remain in the mill it deadens the grinding force by packing the interstices. It diverts what should be grinding energy to the task of re distributing the crop load and of overcoming its frictional resistance and cling. This  though it adds to the heat of milling  contributes little or nothing to the comminution of the larger particles. Where a fairly coarse final grind is wanted (in  say  the 20 mesh zone)  screening affords positive separation since at this size a lively open load can be worked over screens without  blinding . (Blinding connotes the wedging of an oversized particle into a screen aperture  thus putting it out of service  and it becomes serious with clinging material on small screens.) This coarseness calls for strong air currents to sweep finished material out of the mill  and a mill design which includes retaining grates followed by lifters to the discharge overflow may be needed.



The general term  tumbling mill  includes the rod mill  pebble mill  and ball mill. It is of cylindrical or cylindrical conical shape  and rotates about a horizontal axis. A load of crushing bodies called the grinding media forms part of the crop load. They bear upon the piece of ore in the tumbling mixture with abrasive and/or impacting force sufficient to reduce the mineral to particles of the desired size. As in the case of dry crushing  grinding may be divided into two stages  primary and secondary  if the scale of operation justifies such elabora­tion. Milling speeds (r.p.m. and rate of feed)  types of liner  and size and shape of crushing bodies are chosen to develop shattering or impact milling in the primary mill and a more gentle abrasive action in the secondary one. The object is to bring the ore to the mesh of grind called for by the concen­trating section of the mill  and this is best done in separately controlled stages. Often the primary grinding mills must work vigorously on a quickly passing stream of ore  and the secondaries more gently and with longer retention of the pulp.

Types of Mill

Tumbling mills may be classified according to shape into two types  cylindrical and cylindro conical. In the latter two cones are joined by a cylindrical section. At the feed end is a flat cone. After passing the zone of maximum diameter  sometimes called the  drum   the pulp climbs a steep cone to the discharge. This shape has been adopted in the Hardinge mill to develop specialised grinding forces at each stage of the passage of the feed  suited to its changing condition as it progressively disintegrates. Another and perhaps preferable classification of mills is into two types high dis­charge and low discharge. From Figs. 2. and 3. it will be seen that although the feed enters through a hollow trunnion at the centre of the feed end  the discharge arrangement s are very different in the two types. In the high discharge mill the only possible down gradient is a larger trunnion at the discharge than at the feed end  so that in effect pulp only leaves the mill because it has been displaced by entering feed. In the low discharge mill the support of a hollow trunnion at the discharge effect is either avoided in design  or lifting scoops are used  so that a gradient through the mill is created from feed to discharge. In addition to displacement of discharge by entering feed  this adds the effect of gradient to accelerate passage of pulp through the mill at a more or less controlled speed. Since grinding force is applied at an even rate  the faster the ore passes through the mill the coarser is the dis­charged product.

In the type illustrated in Fig.3. the feed  introduced through the scoop at is retained in the crop by the grating C. Material sufficiently fine to pass through this grating is elevated by the radial lifters to the overflow trunnion B. In another  and heavier  form of construction  the discharge end is not hung on a trunnion  but supported on rollers by means of a steel type which encircles the mill shell. The mill discharge can then flow straight out.

It is usual to say that the Hardinge is high discharge only  but this state­ment fails to take account of the effect of the steep cone. The effluent of a mill consists of ore and water flowing as pulp. While it is true that this pulp must climb to the discharge trunnion of the Hardinge  it is also the fact that the cross section of the flowing stream is continuously shrinking  so that the pulp stream must run faster and faster as it flows toward the

One type of mill little used today is the Krupp screen faced cylindrical mill with peripheral discharge. In this machine the feed is introduced through a trunnion and discharged when it has been ground sufficiently to pass through the fine screens lining the cylindrical part of the mill. Heavy perforated plates protect these screens from injury  and a coarse screen is mounted concentrically inside each fine one to give further protection. External sprays provide water and the external casing can be flooded so that the mill dips into water.

The Hardinge Mill

The shell of this mill consists of a flat cone followed by a drum  with a steep cone at the discharge end. The shell is carried in two hollow trunnion bearings  which permit feed and discharge. Drive is by a crown wheel bolted round the steep cone and driven by a pinion. The gear may be straight  single helix  or double helix. Straight or double helix are to be preferred  since no end thrust is set up in operation. Feed may be introduced direct into the feed end trunnion or  more usually  through a feed scoop which gathers ore from the bottom of a feed launder and elevates it to the entrance level. The latter arrangement is particularly useful when the mill is in closed circuit with a  mechanical  classifier. This is a power driven appliance which sorts ore particles discharged from the mill  returning the coarse ones (oversize) for further grinding and allowing the finer ones to overflow as a pulp to the next section. To effect this return (close circuit the oversize)  the returning sands must gain height. They are raked up slope and discharged so as to gravitate to the gathering scoop  which lifts them  together with new feed  to the entry trunnion. This trunnion may have a wearing plate with a conveying spiral to force the feed forward into the body of the mill.

A retaining grid can be used at the discharge end  permitting the mill to be loaded nearly to the centre line without risk of discharging balls. The interior of the shell is protected from direct contact with ore by cast iron or steel lining plates  called  liners . For the cones  segmental sections are used  and for the drum portion curving rectangular liners. They may be backed with plastic material such as rubber sheet or old belting  or fitted directly on to the shell. The liners are held in place by liner bolts kept tight by external nuts and rendered leak proof by plastic washers.

These mills are listed according to the diameter and width of the drum. A small Hardinge has a drum 2  in diameter by 8  wide  weighs over half a ton and when loaded to half volume carries over 4 ton of steel balls. A large mill is 12  by 72   weighs 62 tons  including 27 tons for its set of unworn liners  and carries a ball load of over 50 tons. A 2 h.p. motor would drive the small mill  but between 700 and 800 h.p. are needed for the large one. The power needed per ton of ball load rises with drum diameter  and is some­where between 9 and 15 h.p. per ton of balls at 75% critical speed.

The critical speed of a mill isin r.p.m. when d is the mill diameter in feet less the diameter of the largest ball in the crop load.

Critical speed being a function of peripheral speed  the rate at which these mills can be run depends on the maximum diameter inside fully worn liners. Centrifugal effect is strongest in the drum section  where the ability of the mill to lift its crop load up the rising side of the mill is therefore highest. At the same time the load inside the steep cone lends to work back down slope so the resultant is a pronounced heaping up of load in the drum and a tapering  off toward the discharge. Another differentiating action is also at work  as in all ball mills. The largest balls tend to work to the top of the crop load with the largest lumps of ore. They are then freest to roll or slide down slope on top of the turning load  where they then are most liable to be caught between the bottom of the sliding mass and the downward moving liners There is thus a tendency for the biggest balls and rock to work to the periphery of the drum section. The makers of this mill claim that the balls segregate themselves somewhat  in such wise that the biggest are in the drum and so disposed that when they rise in the turning load they fly or tumble the furthest distance  while the smallest balls work up toward the discharge. Any such ten­dency is useful  since it causes the biggest balls to work upon the newly entered and therefore biggest pieces of ore  while the smaller balls  with less interstitial space  handle the partly finished material as it works its way out. At low speeds the pressure of the heap piled up in the drum has crushing value  and at high speeds balls break clear and rain hammering blows on the mass of churning metal and rock several feet below  giving impact crushing. It is obviously of value to be able  so to speak  to control one s punches  and to use a heavy weight against a tough piece of ore with some selectivity of target while keeping the light hits and tight jostling for the small stuff which does not need such strong treatment. The effect of the conical shape of the Hardinge mill on peripheral speed and kinetic energy on the crop load at various cross sections is illustrated in Fig. 1.

In the attempt to increase the tendency toward segregation of the largest balls at the extreme radius  the Hardinge Co. also market the Tricone mill in which the drum  instead of being truly cylindrical  has a slight back slope toward the feed end.

The Low discharge Cylindrical Mill

Two types of low discharge mill are available. In both  the feed end is served by a hollow trunnion which supports part of the weight  and through which ore and water can either be fed by gravity or  wormed in  by scoop. At .the far end the problem of permitting pulp to discharge at nearly the full diameter of the mill has been solved by two alternative types of construction. In one  the weight of the discharge end is taken by a tyre mounted on rollers  thus leaving the whole end of the mill free for a bolted on retaining grate or a loosely fitting stationary door  which when closed leaves an annular gap through which pulp can escape once its solid fraction is small enough. In the other type a high discharge trunnion is the supporting member  an internal grating with lifting scoops behind it serving to evacuate everything passing through that grating. The grate aperture is chosen at the size at which it is desired to pass worn balls out of the mill  and varies from ¼  to 1 . Grates are cast with a slight flare outware to avoid  blinding  and are procurable in high grade alloy steel. The size of the mill is limited at present by manufacturing possibility. 10  by 10  mills carrying a 45 ton ball charge and using 800 h.p. are in use. A characteristic feature of the low discharge mill is that the diameter  is made as great as possible while length is either about the same as the diameter  or at most twice as great. This follows logically when it  appreciated that the purpose of low discharge milling is to cut down the dwelling time in the mill. A mill in which diameter and length are approxi­mately equal is called a  square mill  in the United States.

Tube  or High discharge Mills

Any low discharge mill can be converted wholly or partly to high discharge by suitably plugging the outlet end  but a better appreciation of the difference between the types is suggested by Taggart  who states that modern practice tends to apply and confine the name (tube mill) to cylindrical mills with a length diameter ratio greater than 2. They were developed  according to Truscott  from the Cornish  barrel pulveriser  which was used to liberate cassiterite middlings  and from cement grinding mills  and were adapted to the needs of cyanidation practice on the Rand. Before the use  now universal  of closed circuit grinding (in which circulation of the ore through the mill and classifier is continued until the latter permits it to leave the circuit) it was of paramount importance that all the auriferous pyrite should be so finely ground as to expose its burden of gold to the chemicals used for its dissolution. The tube mill was therefore made long. In the modern form these machines are from 5½  to 6½  in diameter  up to 22  long  and are loaded with such crushing bodies as pebbles  mine rock  steel balls  or steel scrap.

The Cascade Mill

This recent addition to the range of wet grinding mills departs radically in shape from those  thus far considered. Like the autogenous dry grinding Aerofall mill it has a high diameter/length ratio in this case 3 to 1. Slightly concave liners give maximum diameter at the centre of the drum. This directs the crop load away from the vertical sides  a process aided by deflectors. The mill is made with diameters ranging from 6  to 36  the latter needing up to 6 000 h.p. Feed and discharge are through two support trunnions  finished pulp being lifted to the latter after being ground small enough to pass through a retaining grate.

The vibrating ball mill is finding increasing use in industry  particularly where an inert grinding atmosphere must be maintained.

Body liners receive heavier punishment than end liners  and wear down about twice as fast. A contributory cause of wear is chemical corrosion  either due to mine water or acid from decomposition of sulphide minerals. Fortunately  it is frequently necessary to add lime in the mill circuit for reasons connected with the concentration processes  and this gives protection to the ironwork of the mill  neutralising any acids present. Although the purpose of grinding is the comminution of ore only  the forces employed also act on lining and contents of the mill  and all are subjected to the grinding action. Liners and crushing bodies are therefore selected for their hardness  toughness  and resistance to such wear. When made of alloy steel  they must not introduce any element into the circuit which can cause trouble in the concentrating section  since the steel abrades during grinding and is carried forward to that section. Consumption of steel and cast iron liners varies  and is usually of the order of 0.03 to 0.3 lb. per ton of new feed.

Observations made by Bond in which an experimental glass ended mill loaded with ½  steel balls but no ore  showed that slipping in the layer of balls next the smooth shell increased with the milling speed  in the 60% to 80% critical range  whether the mill was run wet or dry. This effect in a 12  diameter mill not containing mineral would not necessarily be repeated under normal operating conditions. Abrasive grinding is proportional to slipping of the charge. Bond s observations showed 15% slip between the outermost layer of balls and the smooth liners  and a further amount varying from 5% to 10% with each next layer  radially inward to the fifth or sixth layer where observation was not possible.




The element uranium was discovered In Klaproth in 1789. Metallic uranium was first prepared successfully over a hundred years ago by Peligot  who showed that the substance discovered by Klaproth was the oxide and that reduction to metal was much more difficult that had previously been believed. An  interesting history of uranium metal production has been given by Wilhelm. Uranium and its compounds have been of slight com­mercial interest  however  until the recent use of uranium in nuclear applications.

Uranium is the basic nuclear fuel  since it contains the only naturally occurring fissionable material. In these nuclear applications the ele­ment uranium is of interest because of its nuclear properties  providing energy  fission products  and more fissionable material. Urani­um is not important as a metallic material except insofar as  the metal is a convenient form for use in nuclear reactors because it can be fabricated like other metals and has attrac­tive properties such as high density and thermal conductivity. Special  problems arise in the processing of metallic uranium because of its chemical reactivity  radioactivity with consequent health hazards  and anisotropy. Never­theless  the advanced technology developed for the use of uranium as a unclear fuel has yielded many tons of uranium  in various forms   by means of fairly conventional processing. No other metal has ever had its technology so intensively developed over such a short period of time. A large body of information is now available in the unclassified literature.

Isotopes and Nuclear Reactions

For nuclear applications  consideration must be given to the differences in nuclear reactions of the various uranium isotopes. These differ­ences are important enough to justify large scale separation of the isotopes  primarily to concentrate the fissionable uranium 235.

The properties of natural uranium tabulated below must be considered.

Artificial isotopes with mass numbers from 228 to 239 have been prepared. The most important of these is uranium 233  prepared from the reaction of neutrons with thorium 232 and amenable to fission (f 52o).

The plutonium and uranium 23.3 are separated from their respective precursors  uranium and thorium  by chemical means  since  different chemical elements are involved. The separation of the fissionable uranium 235 from natural uranium requires other methods. The principal method for enrichment of uranium 235 is gaseous diffusion  originally carried out on a large scale at Oak Ridge  Tennessee  and now also at Paducah  Kentucky  and Portsmouth  Ohio. These three  plants involve a capital investment of nearly $3 billion and consume 7 per cent of the electrical power generated in this country.


Abundance data for uranium have to be revised upward  as a result of new discoveries stimulated by the incentives for more explora­tion.  Uranium geology  exploration  and devel­opment through the world are still very young  in spite of the tremendous effort that has been put into them in the past few years.  The data on uranium s concentration in the earth s crust  2 to 4 ppm  show that it is about as abundant as beryllium  arsenic  molybdenum  and tantalum. Uranium is more abundant than gold  platinum  silver  cadmium  bismuth  and mercury. Uranium is widely distributed ideologically and also is dispersed  being less concentrated in ores than are other elements of comparable abundance. The bulk of uranium in the earth s crust is believed to be concentrated in a narrow surface zone.

Detailed listings of various uranium minerals are available. Information on the minerals of major economic significance is sum­marized in Table 1. Some important uranium deposits are classified in Table 2.

Originally  the United States obtained its uranium from foreign sources  notably the Bel­gian Congo and Canada. In recent years  domestic sources have been developed to the point where the United States is probably the leading producer of the Western world. Another important  source is the slime residue from gold production in the Union of South Africa (Witwatersrand).

Uranium can also be recovered from other domestic  materials such as phosphate rock  lignite  shale  and monazite sands. Recovery processes have been developed for these marginal sources but are not being applied. With the development of various sources of uranium  notably in the Colorado Plateau  the availability of uranium is not a hunting factor in the devel­opment of nuclear reactors.


In prewar days  before the development or uranium as a source of energy or fissionable material  consumption of uranium compounds was discussed in pounds. This unit is still ap­plicable in the discussion of nonenergy uses of this element. The market provided by these nonenergy uses was too low to justify mining operations primarily for uranium.

Several figures pertaining to the cost of uranium are worth noting. The Atomic Energy Commission has guaranteed a price of U3O5 concentrate. Its price for normal uranium metal billets of normal isotopic con­sent (0.7115 per cent uranium 235) is 840/ kg (818.18. Ib). Enriched uranium is distributed as the liquefied gas uranium hexafluoride (UF4 in steel cylinders. The price per total uranium content or per uranium 235 content is a func­tion of the enrichment  i.e.  the weight fraction of uranium 235. Several examples of the prices are given in Table 4. It is apparent that the cost of highly enriched uranium metal ap­proaches 810 00/lb.

Processing of ores with minerals such as earnotite  containing a high proportion of vanadium  continues to provide a source of vanadium  for which such ores were primarily processed originally.

The wide variety of uranium ores necessi­tates a diversity of mill flowsheets  adapted to the particular ore. The processing steps and the options available to the mill operator are sum­marized in Table 5. A simplified flow scheme is represented in Figure 1. Detailed informa­tion on the various steps is given in an authori­tative work. This volume also includes comprehensive descriptions of mills operating in the principal uranium ore bearing regions of the United States  Canada  and South Africa.

Because of the small quantity of uranium in the. ore  concentration is generally carried out near the source of the ore to minimize ship­ping costs. The capacity of a mill is from several hundred to several thousand tons of ore/day. Convenient tabulations of domestic mills  almost all privately owned  are available.


The uranium concentrate has to be refined to remove the remaining impurities and yield a pure oxide suitable for subsequent conversion through UF4 to metal for reactor fuel or to UF6 for an isotope separation plant. The concentrate is shipped to a feed material processing plant  where both the refining and final conversion to metal are carried out. Fairly conventional chemical operations are employed in the refin­ing  but  since neutron absorbing impurities are harmful even at the level of parts per million  the quality standards are more demanding than those usually encountered  in metallurgical processing.

The concentrate is first dissolved in nitric acid to obtain a solution of uranyl nitrate  UO2(NO3)2. This compound is soluble in various organic solvents such as diethyl ether  methyl isobutyl ketone (hexone)  and tributyl phosphate (TBP). These solvents can be used to extract the uranyl nitrate  which is then re extracted into water. The aqueous solu­tion can be concentrated and evapourated to uranyl nitrate hexahydrate (UNH)  which is  calcined to uranium trioxide  UO3. This oxide can also be obtained by calcination of (NH4)2 U2O7 precipitated from the uranyl nitrate solu­tion. An alternate sequence involves direct pre­cipitation of UO4 with hydrogen peroxide. The UO3 or UO4 is then reduced by hydrogen or ammonia to (brown oxide) UO2. This UO2 may be the form in which uranium is used as fuel. Otherwise  it can be treated with hydrogen fluoride to obtain UF4 (green salt). An alter­native procedure involves reaction of UO2 with NH4HF2 to obtain NH4UF2  which is decomposed to UF4 and NH4F2 which can be recycled. The UF4 is the uranium compound used for reduction to metal. It may alternatively be treated with fluorine to obtain UF6 to be fed to a gaseous diffusion plant where the uranium isotopes are separated.

Feed material processing plants employing the above sequence involving purification of uranium nitrate and generation of pure UO3 are operated by Mallinekrodt Chemical Works at St. Louis and Weldon Springs  Missouri  at Fernald  Ohio. General Chemical Division of Allied Chemical Corp. is using  at Metropolis  Illinois  a process developed at Argonne National Laboratory for more direct preparation of UF6 from the concentrate without going through nitrate and UO3. The sequence of operations applied to the concen­trate resembles that ordinarily applied to refined UO3  obtained from purified uranyl nitrate  reduction to UO2  hydrofluorination to U4  and fluorination to UF6. Fluidization effects reaction between the solid and the gas in each step. Refining is achieved by fractional distillation of the crude UF6 It is also possible to precipi­tate UF4  directly from in aqueous solution by catalytic reduction.

Processes involving nitric acid and uranyl nitrate are used in scrap recovery  carried out mainly at Fernald and in the aqueous process­ing of irradiated fuel.


Pure uranium metal is difficult to prepare because of. the element affinity for other elements such as oxygen  halogens  nitrogen  and carbon. Drastic means of reduction are needed to obtain the metal from stable compounds such as oxides and halides. The reduction has to be performed in closed systems to avoid atmos­pheric contamination. Some of the problems involved in various reduction schemes are better understood with the help of tabulations of boil­ing points of the reactants  melting points of the products  and free energy and enthalpy changes for the reactions.

The large negative free energy of formation of UO2 [–123 kcal/g atom of oxygen at 25°C (77°F)] shows the need for strong reducing agents if UO2 is to be used as starting material in the preparation of the metal. Hydrogen would require a. very high H2/H2O ratio in the gas and cannot be considered a practical reducing agent. Reduction with carbon requires vacuum and leads to contamination by carbide. Calcium is the most favourable reducing agent thermo dynamically  but the heat generation is still so low as to render difficult the separation of the metal and the lime by product. The resulting metal is therefore rather finely subdivided (globules or powder). A halide flux may be added to improve lime removal and thereby obtain coarser metal. Calcium hydride can also be used for the reduction. The metals magnesium  sodium  and potassium are so volatile as to distill from the reaction zone. Aluminium can reduce oxides but  is likely to form alloys with uranium  Since uranium aluminium alloys are used for nuclear fuels  reduction with excess aluminium offers a means of obtaining such alloys directly.

Halides are more suitable than oxides as starting materials for metal preparation. Addi­tional heat is evolved  and the halide by product has a lower melting point . This halide therefore melts and permits the dense uranium metal to settle. Massive metal (biscuit or derby) is thus recovered ill relatively high purity with low losses. In practice  UF4 is preferred over the more hygroscopic UCl4 as the starting material  Na2UCI6 has been used because it is less hygro­scopic than UCl4. In this country  magnesium is the standard reducing agent  being obtain­able in high purity at low cost. It is also being used in England now. The Ames process  developed by Spedding  Wilhelm  and co workers at lowa State College  the use of a sealed bomb because Magnesium s high vapour pressure at temperatures attained during the reaction. The steel bomb  15 in. in diameter by 40 to 45 in. high for over 200 Ib of uranium  is provided with a 1 in. refractory liner  thin enough to permit influx of heat during the preheating period that precedes the reaction  yet thick enough to prevent overheating of the steel by the heat of reaction. Orig­inally  electrically fused dolomitie lime (MgO  CaO) was used for the liner. It has been re­placed by magnesium fluoride recovered as by product of the reduction. Reduction can be carried out on a large scale (e.g.  1½ tons) to give large castings (dingot   direct ingot) not requiring remelting before further fabrication.

In Europe calcium is the preferred reductant for commercial production of uranium via halides. More heat is evolved than with mag­nesium  and the volatility of calcium is low enough to permit reduction at. atmospheric pressure. Calcium was used before magnesium in this country  at Ames and at the National Burean of Standards. Calcium is also used in the United States with an iodine booster for enriched metal  when high yield is an over­riding consideration. Potassium was used in the first preparation of metallic uranium by Peligot  and sodium can also be used. Others have applied these reductants on a laboratory scale  but they hardly lend themselves to large scale operation because of their high vapour pressures.

Fused salt electrolysis at Westinghouse pro­vided the first uranium metal used at the Metallurgical Laboratory.  (No evidence is available of feasibility of electrodeposition from an aqueous solution. Organic solvents  have been studied with inconclusive results.) The Westinghouse process involves electrolysis at 900°C (1652°F) of KUF5 or UF4 dissolved in molten 80 20 CaCI2 NaCl. The metal is deposited as powder on a molybdenum cathode and has to be leached to remove adhering electrolyte. This process provided 65 tons of metal by the fall of 1943  when it was superseded by the Ames process. Bomb reduction has prevailed for commercial production  and electrolytic proc­esses  with various modifications.  are  used only for special purposes. An electrolyte based on UCl3 or UF4 in LiCl KCl eutectic has been developed at Argonne National Laboratory for the electrorefining of high purity uranium near 400°C (752°F).

Two other methods of obtaining uranium in special form are worth noting. The van Arkel deBoer method involves thermal decomposition of a halide  usually iodide  on a hot filament and has been applied to other refractory metals. This method has been applied to the preparation of high purity uranium. When fine uranium powder is needed  the reversible decomposition of UH3 is a convenient source. Reduction of oxide by calcium or magnesium may be used to obtain nonpyrophoric powder directly.


The physical properties of uranium are summarized in Tables 6 to 9. The sensitivity of these properties to purity and metallurgical history accounts  in part  for discrepancies be­tween values reported by different workers. The anisotropy of uranium always has to be borne in mind. The values listed here are based on the critical evaluations by Holden and Klein. References lo the original work may be found in these compilations.  which also in­clude graphical representation and more de­tailed tabulation of those properties as a function of temperature. These compilations have also provided the information in the next section on mechanical behaviour.




Like a number of rare elements  lithium achieved recognition of its potential importance beginning with World War I  at which time Germany used lithium metal for two principal purposes  (1) hardened lead alloy   B Metal   for railway bearings as a substitute for lead tin antimony alloys  and (2) light  strong aluminium alloys   Scleron   in which zinc was largely substituted for copper.

In World War II  interest in lithium grew rapidly for the United States defense developments. Major uses projected for lithium and its compounds in 1954 were as follows

 Anhydrous lithium hydroxide for carbon dioxide absorption.

  Lithium chloride for dry batteries.

 Lithium hydride for hydrogen generation for air sea rescue equipment.

 Lithium hydride as a reagent to produce borohydrides for jet propulsion uses.

 Lithium metal for new alloys such as the magnesium lithium alloys.

 Lithium metal  lithium aluminium hydride  and lithium amide for organic syntheses.

Now  in 1960  the major projected uses for lithium appear to be in the following areas.

Lithium hydride as a reagent to produce borohydrides for jet propulsion uses.

Lithium perchlorate as an oxidizer for solid propellant rockets.

Lithium metal for new alloys such as magnesium lithium and aluminium lithium alloys.

Lithium alkyls as stereospecific polymerization catalysts.

Lithium metal  lithium aluminium hydride  and lithium amide for organic syntheses.

Lithium 6 in atomic energy.

Lithium fluoride as a cell additive in the electrolytic production of aluminium.

It must be recognized that lithium compounds had been used for over a century  before  the World Wars  for ceramics and medicinals  and for some years in the Edison alkaline storage batlery.

The element lithium was discovered by Arfvedsen in Sweden in 1807  during his examination of specimens of the mineral  petalite a lithium aluminium silicate. This discovery was followed by the researches of many scientists  notably Sir Humphry Davy  Gmelin  Berzelius  Bunsen  and Matthiesen.

Lithium is element 3 in the periodic system  only hydrogen and helium preceeding it. It is the lightest metallic element. It has two isotopes  lithium 6 and lithium 7  with an atomic weight of 6.94  indicating a predominance of lithium 7.


Lithium is found widely distributed in the earth s crust  calculated to be approximately 0.004 per cent. By way of comparison  0.002 per cent represents the occurrence of lead  and 0.0006 the occurrence of tin. Lithium occurs as a silicate  phosphate  fluoride  and chloride. The silicate minerals provide the main commercial source of lithium raw material.

Notwithstanding the relatively high quantity of lithium in the earth s crust  actual deposits of sufficient concentration and quantity to warrant commercial mining and recovery operations are limited.

Lithium minerals are found in some quantities on all the continental land masses. From all data available the occurrences would be in the following order of commercial importance  North America  Africa  South America  Europe  Australia  Asia.

If dependent solely on the recovery of high grade minerals  a lithium industry of any magnitude could not be developed. In earlier years the Etta mine in the Black Hills  South Dakota  yielded large quantities of spodumene of high quality  but in that period the demand for lithium products was quite limited. Desert  lake brines in California yield substantial quantities of lithium as a phosphate (Li2NaPO4) but only as a by product of potash and borax production.

The principal lithium minerals useful for commercial recovery are as follows

Spodumene  Li2O.Al2O3.SiO2 containing approximately 8.0 per cent Li2O.

Lepidolite  lithium mica R3Al(SiO2)3 (+F) containing 3 to 5 per cent Li2O.

Amblygonite  LiAl F PO4 containing 8 to 10 per cent Li2O.

Petalite  a lithium aluminium silicate containing 2 to 4 per cent Li2O.

Desert lake brines containing lithium sodium phosphate (Li2NaPO4).

Lithium minerals are usually found in pegmatite dikes which must be of sufficient width to justify mass mining operations. Depth of such mineralization increases with the length of the dikes. Depths usually vary from 200 to 500 ft  in some cases to 1 000 ft. The lithium content of these dikes useful for commercial mining operations is usually from 1 to 2 per cent lithium oxide.

Lithium minerals were usually recovered by hand picking. As the demand increased  mass mining methods were introduced  and the mineral concentrates were produced by two methods  (1) froth flotation  and (2) heavy media sink float (magnetic reagent). Froth flotation appears to be the most satisfactory method to produce concentrates containing 4 to per cent Li2O.

Cost Considerations

As in other metallurgical products  the cost of lithium metal is largely dependent on the  cost of raw material. A minimum size of mine flotation mill for economic recovery of lithium is 150 to 200 tons raw material/24 hr  producing 30 to 40 tons of 4 to G per cent concentrates. In 1946  such concentrates could be produced for $4/unit of lithium oxide at the mines. In 1952  these costs were of the order of $7/unit of lithium oxide  which is equivalent to 17.5 cents/lb of lithium carbonate recovered (80 per cent recovery).

To produce 1 000 lb of lithium metal  a minimum of 6 250 lb of lithium chloride is required. To produce this amount of lithium chloride  5 700 lb of lithium carbonate are used  or 2 280 lb of lithium oxide. If extraction recovery is maintained at 90 per cent  then approximately 25 tons of raw material concentrates containing 5 per cent lithium oxide would be required. Therefore the raw material cost per 1 000 lb of lithium metal would be as shown above.

Allowing for the cost of extraction  salt conversion  and electrolysis to produce metal  and also for capital costs  overhead  and profit  the price of lithium metal in quantities of 100 000 Ib/annum or more will be approximately 10 to 15 times the raw material cost.

Capital costs including working capital will be of the order of $5 to $8/lb/annum production including all operations from extraction to metal production.


Having obtained ore concentrates  the problem of extracting and recovering lithium becomes all important.

Methods tried and utilized are as follows

Heating with potassium bisulfate and then leaching out lithium sulfate.

Using spodumene to furnish alumina and silica to produce portland cement. Spodumene is added to limestone and calcium chloride. The entire mixture is calcined in a kiln to produce cement clinker  and the lithium is volatilized as an impure chloride.

Heating spodumene concentrates at 1000°C+ (1832°F) (below fusion) to convert from alpha phase to beta phase spodumene. The material is cooled and finely ground  incorporated with sulfuric acid  and heated to approximately 300°C (572°F)  then cooled and leached to produce an impure solution of lithium sulfate.

Heating spodumene ore at 1038°C (1900°F) or lepidolite ore at 871°C (1600°F) with 3 parts of ground limestone in a rotary kiln. The sinter is leached with water to give lithium hydroxide monohydrate.

The First method was utilized in Germany but did not give maximum yields or lowest costs. It was primarily a batch process.

The second method affords an interesting means of extracting lithium as a by product  provided it is carried out on a large scale wherein portland cement can be commercially produced at least 1 000 bbl/24 hr. Ninety eight per cent of the lithium can be volatilized  but the subsequent purification of the condensed lithium chloride presents very difficult problems.

The third method offers a continuous process with good economy whereby at least 90 per cent of the contained lithium can be recovered. The sulfate solution is easily filtered to remove the aluminium silicate residue  purified of minor impurities  and recovered as quite pure lithium carbonate. Soda ash (Na2CO3) is added to the sulfate solution precipitating insoluble lithium carbonate with Glauber s salt (Na2SO4.10H2O) as a by product.

The fourth method offers recovery percentages lower than the third method. When treating spodumene  the inversion step is not necessary.

Table 1. summarizes  the processing operations of the various United States producers of lithium compounds. The third method described above is the Lithium Corp. of America process  and the fourth method is that practiced by American Lithium Chemicals and Foote Mineral Company.

With pure lithium carbonate or hydroxide as a starting chemical  all of the useful lithium compounds can be produced advantageously.

Other methods of extraction and recovery have been considered  among them the volatilization of lithium compound from electric furnace reductions. Lithium minerals and compounds are excellent fluxes  but as such are quite costly  necessitating very complete volatilization and recovery of the lithium  or absorption in slag with subsequent costly method of extraction and recovery.

Production of Lithium Metal by Fused Salt Electrolysis

In this chapter attention is focused primarily on lithium metal. Various methods have been proposed to reduce lithium compounds to metal  but results are costly as compared to the classical method. Lithium chloride admixed with potassium chloride is fused. Current is applied  depositing lithium metal on the cathode and evolving chlorine at the anode. In many respects the technology is comparable to the electrolytic reduction of magnesium chloride. Some typical operating data for commercial cells are given in Table 2.

Unit cell capacity is relatively small 80 to 100 Ib of lithium metal/24 hr. When demand would justify it  cell capacities may be increased advantageously to possibly 500 Ib  24 hr. For large scale production  cells may be operated in series to permit the use of standard electrical equipment  generating 125 to 250 volts d c.

Using salts approximating a euteetic mixture  the temperature of the molten bath can be maintained at a minimum 450 500ºC (842 932°F) (LiCl 60 per cent  KCl 40 per cent approximately). Deterioration of the graphite anodes is thereby minimized.

Lithium metal thus produced will contain certain impurities  principally sodium. Practically all of the sodium present in the lithium chloride is deposited with and remains in the lithium metal produced.




In 1797 Vauquelin discovered the element  beryllium as a constituent of the mineral beryl. In the French language the element is referred to as glucinium (Gl). This name is derived from the sweetish taste of many of its compounds.

The first metallic beryllium was produced by Wohler and Bussy in 1828. They obtained beryllium in the form of an impure powder by reducing beryllium chloride with metallic potassium.

During the nineteenth century numerous other investigators contributed to the development of the chemistry of beryllium. Of particular interest is the work of the French scientist Lebeau  published in 1899  which includes descriptions of the electrolysis of sodium beryllium fluoride resulting in the production of small  hexagonal beryllium crystals  and the preparation of beryllium copper alloys by direct reduction of beryllium oxide with carbon in the presence of copper. Also of interest is the work by the German scientist Oesterheld  who  in 1916  published the equilibrium diagrams of beryllium with copper  aluminium  silver  and iron  and the investigations by Wilhelm Kroll  W. B. Donahue  and C. Adomali  who all worked on methods of producing pure beryllium by reducing mixtures of beryllium fluoride and alkaline earth and alkali fluorides with alkaline earth metals and magnesium.

Commercial development of beryllium in the United States was begun in 1916 by Hugh S. Cooper  who produced the first significant beryllium metal ingot  and by the Brush Laboratories Company  which started their development work under the direction of C. B. Sawyer in the early 1920 s. In Germany the Siemens Hakke Konzern began their commercial development work in 1923.

Beryllium alloys made their commercial appearance in the United States in 1932  when the American Brass Company  using beryllium copper master alloy produced by the Beryllium Corporation of America  made available the first rolled beryllium copper. The developments of the Brush Laboratories Company were taken over by the Brush Beryllium Company in 1931. This company entered the beryllium copper field in the early 1930 s and was the first company to develop a commercial process to produce and market metallic beryllium in solid form.

Beryllium is the first element in Group II of the periodic system of elements. Its atomic number is 4  its atomic weight 9.013  and it has a valence of two corresponding to 2s electrons in the L shell. No isotopes have been found in beryllium occurring in nature. Radioactive isotopes of mass numbers 6 7  8  10  and 11  however  have been artificially produced.

Metallic beryllium is grayish in colour. Large crystals of bright metallic luster are usually discernible. It is a very light metal (specific gravity 1.845) and is known as the only such metal combining good physical strength with a high melting point [1285°C (2345°F)].


There are some 30 recognized minerals containing beryllium  only 3 are of significance  viz.  beryl (3BeO Al2O3 6SiO2)  phenacite (2BeO.SiO2)  and bertrandite (4BeO 2SiO2H2O). Of  these  three   only  beryl  is now  of industrial   importance.   However    substantial low grade deposits of the other two minerals have now been found. The recovery of beryllium from these low grade ores is now under investigation. In pure form this mineral is a beryllium aluminium silicate containing approximately 14 per cent beryllium oxide (BeO)  19 per rent  aluminium oxide (Al2O3)  and 67. per cent silicon dioxide (Si02). The pure composition  is  approached  in the  precious  forms of beryl emerald and aquamarine.

Industrial grades of beryl ore now reaching the market contain approximately 10 to 12 per cent beryllium oxide. Rarely is ore of more than 12 per cont beryllium oxide available  and the trend is toward a supply ranging from 11 percent downward to the marginal ores containing less than 9 per cent. Other constituents of the ore are aluminium  oxide  17  to  9  per cent  silicon dioxide  64 to 70 per cent  alkali metal oxides  1 to 2 per cent  iron  1 to 2 per cent  and minor amounts of other oxides. Feldspar  quartz  and mica are the principal mineral contaminants of   commercial grades of beryl ore.

Occurrences of beryllium in the earth s crust are  widely  distributed  and  are  estimated  to amount to approximately 0.001 per cent. Beryl ore  containing  10  to   12  per cent beryllium oxide  however  has not as yet  with one or two exceptions  been found anywhere concentrated in large enough  quantities to be mined economically for its own sake. The  supply  is  therefore  generally obtained as a by product of mining feldspar  lithium  or mica in pegmatite  dikes  and only those crystals which are large enough to be hand sorted and cobbed are recovered. The best producing pegmatites  suitable for hand sorting and cobbing  contain 1 to 2 per cent beryl ore. The by products from these operations and low grade beryl ore containing about 0.1 per cent beryl will require beneficiation  processes  to   recover  the  beryl. According to the Bureau of Mines  reserves of such  low grade   ores in the United States amount to over 1 000 000 tons of beryl ore content.

Recently  privately owned facilities for concentrating low grade beryl ores have been constructed in Colorado. The feasibility of operating these mills on an economic basis has not yet been established.

The principal producers of beryl ore are the Union of South Africa  Argentina  Brazil  and India. Small amounts are produced in British East   Africa  French Morocco  Mozambique. Portugal  and Canada. In the  United States the principal sources are found in Colorado  Maine   New  Hampshire  and South Dakota. Many undeveloped deposits are located in Canada  particularly in the Great Slave Lake area of the Northwest Territories. Most of the beryl ore consumed by the beryllium industry in the United States is imported. Up to this time  the supply of hand sorted and cobbed ore has been more than sufficient to meet the industrial demand.




Pure Beryllium Oxide

The technical grades of beryllium hydroxide or beryllium oxide produced as intermediate products in the production of beryllium metal and beryllium copper master alloy contain  as a major impurity  salts of sodium and also minor amounts of elements occurring in commercial grades of beryl ore (Si  Fe  Al  Cu  Mg  Li  Mn  etc.).

These impurities are generally removed or substantially reduced by dissolving the technical grade hydroxide in sulfuric acid and purifying the formed beryllium sulfate by re crystallization. The pure beryllium sulfate crystals obtained in this manner are then heated in gas fired firebrick furnaces to drive off sulfur trioxide (as SO2)  leaving a pure grade of beryllium oxide as a residue. To remove all sulfur in the oxide  temperatures of 1000  1200°C (1832 2192°F) are required in a reducing atmosphere.

An extremely pure but very expensive grade of beryllium oxide can also be produced by converting the technical grade beryllium hydroxide into beryllium base acetate [BeO.3Be (C2H302)2]. The impurities remaining in this salt are separated by extracting the basic acetate with chloroform  filtering the solution to remove the insoluble impurities  evapourating the chloroform  and subliming the residue. The pure basic acetate so obtained is subsequently fumed down with chemically pure sulfuric acid. The resulting pure beryllium sulfate is then decomposed as noted above.

 Beryllium Metal

All pure beryllium metal now produced in the United States is made by the reduction of beryllium fluoride with magnesium metal. The beryllium fluoride used is produced from beryllium hydroxide.

In this process  developed by Kjellgren and used by The Brush Beryllium Company  an excess of beryllium fluoride is used in relation to the amount of magnesium added.

The reaction  which is carried out in a graphite lined furnace  is kept under control by gradually charging magnesium metal and beryllium fluride in solid form at a furnace temperature of about 9000C (16520F). The head generated by the reaction is absorbed as fast is it is liberated and serves to supply part of the heat required to melt the solid magnesium and the beryllium fluoride. Since the reaction is carried out below the melting point of beryllium  the metal is produced as very fine particles dispersed in a slag consisting of the magnesium fluoride formed by the reaction and the excess beryllium fluoride used.

The beryllium particles are melted and coaleseed by raising the furnace temperature to somewhat above the melting point of beryIlium. At this temperature the slag is very liquid  because of the presence of the excess beryllium fluoride  and the molten beryllium separates well and floats on the molten slag. As soon as this occurs  the molten metal and the slag are poured out together into a cold graphite crucible and solidified. In the process of pouring  the molten beryllium metal is broken up into pebbles so that the solid mixture obtained in the graphite crucible consists of beryllium pebbles embedded in solid slag. This mixture is broken up  and the beryllium pebbles and the excess beryllium fluoride present in the slag are recovered in the process which is used for producing the beryllium fluoride from beryllium oxide.

In this process the beryllium oxide is dissolved in a solution of ammonium hydrogen fluoride. The ammonium beryllium fluoride solution produced in this manner  and the ammonium beryllium fluoride solution recovered from leaching out the excess beryllium fluoride present in the mixture of the pebbles and slag resulting from the reduction  are mixed together and treated with lead peroxide  calcium carbonate  and ammonium polysulfide to precipitate impurities. After filtration  extremely pure ammonium beryllium fluoride is recovered from the filtrate by evapouration and crystallization.

In this manner the pure ammonium beryllium fluoride crystals produced in the process are derived both from beryllium oxide  the raw material  and from the beryllium fluoride present in the recycled slag.

The crystals obtained are then decomposed by heat into ammonium fluoride vapour  which is absorbed in water and recycled to the step of leaching the beryllium fluoride containing slag  and into molten beryllium fluoride  which is collected in the form of small clear droplets or lumps.

After leaching the mixture of pebbles  and slag resulting from the metal reduction step in order to recover the soluble beryllium fluoride  the beryllium pebbles are easily separated from the remaining insoluble  small crystals of magnesium fluoride.

BeryIlium pebbles obtained in this manner are remelted in a beryllium oxide crucible heated in a vacuum furnace.

The purity of the metal is controlled by the purity of the beryllium fluoride and the purity of the magnesium used.

A flow sheet of the Brush Beryllium Company process for recovering beryllium metal from beryllium hydroxide is shown in Figure 3.

Processes for producing beryllium in flake form by electrolysis of molten mixtures of beryllium and alkali chlorides have been worked out by Cooper  Sawyer and Kjellgren  and Morana. These processes have  so far  not been used in industrial practice in the United States. Thepechiney Company in France has  however  over a period of years  produced beryllium flake by chloride electrolysis  although it is reported that their production is minor in comparison with the production in the United Slates.



In 1817 John Jacob Berzelius  a professor of chemistry in Stockholm and secretary of the Swedish Academy of Science  was studying  along with J. G. Gahn  a method formerly in use at Gripsholm  Sweden  for the production of sulfuric acid. As a part of this study  an examination of the acid itself revealed a sediment which gave forth an offensive odour previously identified by Kloproth as an indicator of the presence of tellurium. J. G. Gahn  it appears  recalled detecting a similar odour in those plants where the Fahlun copper concentrates were smelted and the resulting sulfur used to produce sulfuric acid. In the hopes of finding a new source of the then rare element  tellurium  in this acid  Berzelius obtained larger quantities of this residue  but his work proved to no avail for he could find no trace whatsoever of tellurium. He did notice  though  as a result of his tests  that there remained unaccounted for an unknown substance whose chemical properties closely resembled those of tellurium. So closely akin were these two elements that Berzelius decided to call the former selenium  from the Greek word selene meaning the moon  tellurium having been derived from the Latin word tellus  meaning the earth.

Although discovered in 1817  selenium remained a laboratory curiosity for some 50 years. Finally  in 1873  Willoughby Smith  while testing various materials for electrical conductivity  discovered quite by accident that the current resistance of this element decreased as the intensity of illumination increased  and furthermore  that resistance increased slightly as the temperature likewise increased above 170°C (338°F). This led  among other things  to the development of the photoelectric cell  of which more will be said later on  but the important fact is that for the first time it brought selenium into the public eye. Once there  a multitude of applications developed  until now plays a very definite part in our everyday life.


Selenium  the fortieth clement in plentifulness  falling between bismuth and gold  rarely occurs in its native state. Although occasionally found in conjunction with native sulfur  and in the form of selenides of other metals in such minerals as clausthalite  PbSe  eucaisite  CuAgSe  crookesite  (CuTIAg)2Sc  naumannite  Ag2Se  and zorgite  PbCuSe(?)  it is most frequently found as an accessory mineral in base metal ores of lead  copper  and nickel. Whenever any of the above are treated  seleniuim is recoverable as a by product.

The main sources of selenium in this country originate in the copper mining states of Utah  Arizona  and New Mexico  the Montana ores running low in this element. Selenium from these sources and from Mexican copper mined in the states of Durango and Zacetacas is recovered by the American Smelting and Refining Company  Anaconda Copper Mining Company  the American Metal Climax Company  and  recently  Kennecott Copper Corporation in this country. In Canada the International Nickel Company of Canada  Ltd.  recovers selenium from the Sudbury copper nickel ores  and the Canadian Copper Refiners  Ltd.  of Montreal from the copper anodes derived from the copper zinc ores of Flin Flom  Manitoba. In Australia selenium is produced by the Electrolytic Refining and Smelting Company  Pty. Ltd.  in Sweden by  the Boliden Mining Company  in Belgium by the Societe General of Hoboken  in Japan by the Taihi  Besshi  and Nippon mining companies  and in the Western Zone of Germany by the Norddeutsche Affinerie of Hamburg. Prior to World War II  selenium was produced in what is known as the East Zone of Germany  but that source  along with all other production in the U.S.S.R. and other Soviet dominated countries behind the Iron Curtain  has become an unknown factor today.


Many years ago the sole source of selenium was thought to be from the flue dusts of metallurgical processes utilizing sulfide ores  however  recovery from this source is virtually nonexistent today  and the anode muds or slimes from electrolytic copper refineries provide the source of most of the world s selenium. Basically  there would appear to be three main methods of recovering selenium by roasting with soda  by roasting with sulfuric acid  and by smelting with soda and niter  variations of these are found to accommodate the variances in basic raw materials being handled.

In the first method  that of roasting with soda  the decopperized slimes are mixed with soda and raised to temperatures well below the sintering point with sufficient air to oxidize the selenium. The selenium is recovered as sodium selenate from the water leached calcine by evapouration. The sodium selenate is reduced to the selenide by coke  then redissolved  and blown with air  and the selenium is precipitated with sulfur dioxide  the sodium hydroxide being carbonated and recirculated.

In the sulfuric acid roasting the raw slimes are treated with sulfuric acid prior to roasting. During the course of roasting  the selenium dioxide is driven off and collected in a wet scrubber cottrell system. The remainder of the selenium is largely  recovered by the conventional soda smelting process.

The soda smelting process is a pyrometallurgical one in which the decopperized slimes are mixed with soda and silica. After the first slags are drawn off  the molten charge is rabbled with air  and some of the selenium is volatilized and caught in a scrubber Cottrell system. To the charge is now added caustic and niter. The slag which results  high in both selenium and tellurium  is crushed and leached with water  to which is added fresh sulfuric acid to precipitate the tellurium. The solution is then treated with sulfur dioxide to precipitate the selenium.

There are a present in the United States and Canada seven producers of this element  with a total output of between 1 and 1.4 million pounds annually. Stimulated by the Korean conflict  increased demand for seIenium  especially for rectifier manufacture  starting in 1951  led to an acute shortage for this element. Material improvements in recoveries by selenium producers resulted in increased production  ending the shortage by the fall of 1956 At present  and for the foreseeable future  the supply of selenium should be adequate to take care of any reasonable demand  based both on production potential and stocks in the hands of producers.

Pricewise  selenium can hardly be termed an inexpensive commodity at $7.00/lb retail in its commercial form  as quoted by domestic producers. This price for domestic material has ranged from a low in the early 1930 s of $1.50/lb to a high of $15.50/lb in 1956.


Selenium  atomic number 34  is the third member of Group VI of the periodic arrangement of elements. Selenium is more metallic than sulfur but less metallic than tellurium  the two next of kin in the group.

Selenium can be caused to exist  as a solid  as a liquid  or as a vapour at temperatures easily handled in any metallurgical laboratory.

Amorphous Selenium. Amorphous selenium occurs as red powder  vitreous  and collodial selenium.

Red Powder. The amorphous red powder results when solutions of selenous acid of pH 7 to pH 3  and even more acidic  with such acids as hydrochloric or sulfuric  are treated with strong reducing agents such as sulfur dioxide  hydrazine  or hydroxylamine hydrochloride. The red powder turns black on standing  and  on heating  yields the hexagonal form.

VITREOUS. Vitreous selenium  a black mass prepared by quench cooling liquid selenium  is. glassy and brittle  showing conchoidal fractures  and is described as a supercooled liquid. It is a dielectric  being electrified by friction.

As little as 0.0003 per cent by weight of tellurium in the selenium will increase the plasticity. Vitreous selenium containing 1 per cent of chlorine is quite plastic when first quench cast  but within 24 hr  even at room temperature  it will show substantial conversion to the hexagonal state.

By reflected light  glassy (vitreous) selenium is mirror black. Thin layers by transmitted light (daylight) appear blood red. Vitreous .selenium is about as hard as glass  is perhaps more brittle  and shows a pronounced conchoidal fracture. Vitreous selenium is a very poor conductor of heat but  if stored in a warm place  gradually crystallizes to the hexagonal state which is a better conductor of both heat and electricity. Except for semiconductor use  especially xerographic plates  vitreous selenium  as such  may be considered an electrical insulator.

COLLOIDAL. Colloidal selenium is prepared by the reduction of dilute aqueous solutions of soluble selenium with such reducing agents as sulfur dioxide  hydraline hydrate  dextrose  or titanium trichloride. It can also be prepared by pouring a solution of selenium in carbon disulfide into a large volume of ether  and by passing an electric current through a solution of selenous acid between a platinum anode and a selenium coated cathode.

The colours obtained vary from violet to red  depending on the conditions during precipitation.

Crystalline Selenium. Crystalline selenium occurs as either monoclinic or hexagonal forms.

MONOCLINIC. Monoclinic selenium is obtained by the low temperature evapouration of carbon disulfide containing dissolved selenium. Heat and several other conditions convert the mono clinic selenium to the hexagonal state.

HEXAGONAL. Hexagonal selenium is considered the most stable state of selenium under ordinary conditions. It has a gray metallic appearance  is a fair conductor of heat and electricity  is fairly inert to atmospheric conditions  has fair mechanical strength  and is easy to produce by heating any form of selenium until crystallization is complete.

The Liquid State

The melting point of hexagonal  selenium is 217°C (423°F)  and represents a definite transition from solid to liquid. Amorphous selenium begins to soften at about 1000C (2120F)  and its behaviour at increasing temperatures depends on the rate of heating. It liquefies at 2170C (4230F).

Molten pure selenium  no water from what allotrope it is obtained  is not very fluid until heated several dozen degrees above the melting point of 2170C (4230F).

Unlike sulfur  selenium becomes more fluid with increasing temperature. Liquid selenium can be boiled at atmospheric pressure [684.90C (12650F)] in a porcelain dish or in a quartz dish without bumping. Liquid selenium probably contains several molecular species.



The first known reference to platina or native platinum  a naturally occurring alloy composed of a large proportion of platinum together with palladium  rhodium  iridium  osmium  rutheni­um  copper  iron  and sometimes gold  bears the date 1557. The writer was Julius Caesar Scalinger (or della Scalla)  an Italian poet and scholar  who noted the difficulty in melting a metal which was obtained from the Spanish possessions in South and Central America. The earliest scientific investigation of platina was instigated by William Brownrigg in 1750. Ber thollet and Pelletier described the work of M. de I Isle  who obtained a malleable form of platinum in 1773 74. Pierre Francois Chebaneau (or Chavaneau) succeeded in preparing some malleable platinum  and patented the process in 1783. However  an intensive study of the records has led to the belief that pure platinum was first obtained in 1803 by W. H. Wollaston  whose brilliant researches resulted in the isolation of two of the minor constituents of platina  i.e.  palladium and rho­dium. The aqua regia extract of platina was treated with ammonium chloride to precipitate the platinum  mercury (I) cyanide was then added to precipitate the palladium  and finally rhodium was isolated as sodium rhodium chloride.

The name  palladium  was chosen by Wollaston in honor of the asteroid Pallas  and  rhodium  because of the rose red colour of the salts of that metal. Osmium and iridium were isolated and named in 1804 by S. Tennant. About the same time  H. V. Collet Descotils  A. F. de Fourcroy  and L. M. Vauquelin also suspected the presence of iridium in platina  but their work was not conclusive. Both osmium and iridium were found in the black residue which remained after the aqua regia treatment of platina. The name  iridium  was derived from the Greek word iris  meaning a rainbow  and refers to the varying colours of iridium salts. The name  osmium   derived from the Greek word for smell or odour  was chosen because of the characteristic chlorine like odour of osmium tetroxide. It was not until 40 years later that the remaining clement of the platinum group was isolated. Ruthenium was discovered and named by  C. Claus in 1844. although publication of the news was delayed until the following year when J. J. Berzelius confirmed the results of the experiments and accepted ruthenium as a new element. The name  ruthenium   from Ruthenia  Russia  was first used by G. Osann in 1828 to designate a substance obtained from platina  this substance was later shown to be composed of the oxides of silicon  zirconium  titanium  iron  and a small quantity of a new element.


The platinum metals occur both in primary deposits and in placers. The primary deposits are of two main types. The first consists of disseminations or local concentrations of the metals in olivene rich rocks  particularly in dunite and often associated with chromite  native platinum  or iridosmine  is the principal constituent. The erosion of such deposits has. been responsible for the formation of placer deposits of the platinum metals. Dunite de­posits are widespread  the most important commercially being in the Ural Mountains region of the U.S.S.R.  and at Overwacht in the Transvaal  Union of South Africa. The second type of primary deposit includes the magmatic nickel copper sulfide deposits which are generally associated with norite. These de­posits  in which platinum and palladium pre­dominate  make up the greatest known reserves of platinum metals. The most extensive deposits have been found in the norite belt of the Bushveld igneous complex in the Transvaal  and in the Sudbury district of Ontario  Canada. In the Sudbury district ores  platinum and pal­ladium occur in about the same proportions  these ores also contain small amounts of the other platinum metals as well as silver and gold. The precious metals are obtained as by products during the extraction of nickel and copper. The South African primary deposits contain all the platinum group metals  as well as iron  nickel  copper  cobalt  silver  and gold. In contrast to the Ontario ores  the quantity of base metals present is not sufficient to pay for the working costs. Platinum metals have also been found in quartz veins and in copper and coal deposits  although these sources are of little economic value.

Placer or alluvial deposits of great economic importance have been found in the Perm dis­trict of the Ural Mountains  in Colombia  South America  and in Abyssinia. Placers have also been found in the United States including Alaska  Australia  and Canada.

Very few compounds of the platinum metals occur as minerals. Sperrylite (PtAs2  with which small amounts of rhodium are sometimes asso­ciated) occurs in the Sudbury district in Can­ada  in the Transvaal deposits of South Africa  and in eastern Siberia. Cooperite (PtS) and braggite (Pt  Pd  Ni)S] are found in the Bushveld complex  and in the Potgietersrust districts of the Transvaal. Of the minerals which contain palladium principally  stibiopal ladinite (Pd3Sb) is found with sperrylite in the Transvaal  and potarite (PdHg) is found only in British Guiana. Laurite (RuS2)  which some­times contains osmium  is very rare  it is found in Borneo and in the Transvaal.

Iridium is most often found alloyed with os­mium in iridosmine (Ir > Os) and siserskite (Os>Ir). The term osmiridium is used more or less synonymously with the term iridosmine  but the latter is employed in this chapter. Iridium is also associated with platinum and gold in platiniridium and aurosmiridium  re­spectively. As mentioned previously  rhodium  together with other members of the platinum group  occurs as a minor constituent of native platinum  the proportion of rhodium is usually less than 1 per cent. Small quantities of ruthe­nium are associated with alloys of the platinum metals  but most frequently with iridosmines which may yield 12 per cent or more of ruthe­nium. It must be noted  however  that some types of iridosmines are very resistant to cor­rosion  and as a result the analysis errors may be large.


In the first edition of this handbook  the authors included estimates of world production of crude platinum up to the end of 1916  and of average annual outputs by countries from 1921 to 1952. This information is largely of historical interest  and has not been retained for this edition. It is perhaps interesting to note that  in the period from 1950 to 1957  world production of the platinum metals doubled. This increase is mainly the result of a greater than fourfold expansion in the South African production  this country s output now accounts for more than half of the estimated world production. In the same period Canadian output increased by approximately 50 per cent  while that of the U.S.S.R. is esti­mated to have increased by 25 per cent. In 1958 a drop of about 30 per cent in world pro­duction was reported.

Notwithstanding the record production of these metals in 1957  figures for world consump­tion showed a decrease from the previous year  and a further decrease of about 20 per cent was estimated for 1958. This has been attrib­uted mainly to a reduction in demand from the petroleum industry in the United States but undoubtedly it is related to the period of economic recession in North America in 1957 58  which saw a decline in consumption of many base metals as well. In 1959 industrial demand increased  and in consequence there was a limited increase in production in both South Africa and Canada.

It is significant that increased production and de­creased consumption brought the prices of these metals down appreciably below the figures for the previous 6 years. In fact  the price of platinum for January 1  1959  was as low as at any time since World War II. With the resump­tion of demand in 1959 there has been a slight upward turn in prices  although these remain lower than the average of the previous five years.


The methods by which platinum metals are obtained from their naturally occurring sources and from scrap precious metals should be dis­cussed under various headings according to the source of the raw material. It may be advisable to point out that  by tradition  refiners of the  platinum metals have long been reluctant to disclose the details of their operations  accord­ingly  the descriptions that follow  which are based on published material only  may depart in detail from prevailing practice.

Extraction of Platinum Metals from

Canadian Nickel Ores

As has been mentioned earlier in this chap­ter  platinum metals occur associated with the copper nickel sulfide ores in the Sudbury dis­trict of Ontario. The total content of platinum metals is only of the order of ½ ppm  but  in view of the large tonnages of nickel produced  Figure 1. Process Flow chart for the recovery and separation of platinum metals. First stage  ore to platinum concentrates the quantities of platinum metals made available are sufficient to make this an important  source. The bulk of the nickel ruining in the Sudbury basin is carried out by the International Nickel Co.  a smaller proportion being mined by tin Falconbridge Nickel Co. Expansion of the pro­duction of the former company includes a new project in the Mystery Lake Moak Lake dis­trict of northern Manitoba  about 400 miles north of Winnipeg. Present indications are that this ore body will also produce platinum metals as by products  and that this  known as the Thompson Mine  will become the second ranking producer of platinum metals in Canada Production is expected to begin in 1960 or 1961. The manner of obtaining the platinum metals from these ores is obviously integrated with the processes employed for the isolation and refining of nickel and copper. These are de­scribed in recent reference works on production metallurgy. The essential stages in the operations of the International  Nickel Co. are outlined in Figure 1. The bulk of the platinum metals is separated from the nickel and copper during slow cooling of the Bessemer matte. During the preparation of this  the amount of oxidation of sulfur is con­trolled to produce a small amount of metallic nickel and copper which acts as a collector to separate the platinum metals from the metallic sulfides. This precious metal alloy is magnetic  and may be removed by passing the ground matte through a magnetic separator. This prod­uct is melted and treated with enough sulfur to convert 80 to 90 per cent of the nickel and copper to sulfides  at the same time retaining a small proportion of these metals in the free state. On cooling this matte a still more con­centrated metallic alloy containing the platinum metals separates  and is removed from the ground material magnetically. This enriched alloy can then be subjected to electrolytic re fining  during which the platinum metals accumulate in the anode slimes.

Falconbridge Nickel Mines  Ltd.  separate a rich concentrate by magnetic cobbing  then subject the remainder to flotation and a further wet magnetic separation. The concentrate is smelted in a blast furnace to give a matte which is further upgraded by treatment in basic lined converters to yield a high grade ship­ping matte. This matte  containing about 48 per cent nickel  28 per cent copper  and 22 per cent sulfur  is shipped to the company s sub­sidiary refinery at Kristiansand  in the south of Norway  for further treatment. There it is roasted  the copper removed from the calcine by acid leaching  and the nickel recovered by electrolysis. The anode slimes from this electrolysis are smelted and re electrolyzed to give a concentrate of precious metals which is refined in the same plant in Norway. The oper­ations of this company have been described in detail in a recent publication.


Tantalum  atomic number 73  atomic weight 180.95  is located in Group VB of the periodic table below its sister element  columbium  and adjoined by hafnium on the left and tungsten on the right.

It is a strong ductile metal characterized by  (1) its high density  16.6 g/cc  (2) its high melting point  2996°C (5425°F)  the third highest among the metals  exceeded only by rhenium. 3180°C (5756°F) and tungsten. 3410°C (6170°F)  (3) the tenacious thin oxide layer on its surface which gives it superior rectifying and capacitance properties  and (4) its extreme inertness to attack by all acids  except hydrofluoric and fuming sulfuric  at ordinary temperatures.


Tantalum was discovered in 1802 by Ekeberg of Sweden  he named it after Tantalus in Greek mythology because of the difficulty of dissolving the oxide. In 1801 Hatchett of England had announced the discovery of columbium  because of the similarity of the properties of the compounds of those two elements  for over forty years the two were regarded as identical  although Wollaston suspected their dissimilarity. In 1844 H. Hose  a German chemist  made exhaustive studies of the columbite of Bodenmais and showed that this mineral contained two metallic acids  one of tantalum and the other of what he supposed to be a new metal which he named niobium (for Niobe  the daughter of Tantalus).

According to the distinguished American chemist J. Lawrence Smith  in a note published in 1877  Rose believed in 1844 that the tantalum of this mineral was the same as what was equally well known as columbium. Thus it appears that he regarded the other metal in the mineral as a new element and did not simply apply another name to Hatchett s columbium.  Subsequent examination  however  convinced Rose that the two metallic acids obtained from the Bodenmais columbite were really the original columbic acid of Hatchett  discovered in 1801  and the tantalic acid discovered by Ekeberg in 1802.  In the face of this it is not understood why Rose s new name  niobium  was accepted then  as well as in more recent years  in place of Hatchett s original and valid name  columbium.

In 1866 Marignac developed his classical method of separating the two sister elements utilizing the difference in the solubilities of their complex potassium fluorides. This permitted studies of the compounds of each element.

None of the early investigators actually isolated anything more than an impure form of either metal. The first ductile tantalum was produced by W. von Bolton in 1903 in the Siemens Halske plant in Berlin. It was the first metallic filament for incandescent lamps  and approximately 11 million tantalum lamps were made before tungsten began to replace tantalum for this application in 1909.

Tantalum was first produced in the United States in 1922 by C. W. Balke and commercial production has continued since  that time. It is interesting to note that one of Balke s motives in making tantalum was to take advantage of its chemical inertness and use it as an anode in the corrosive chlorine cell electrolyte. He soon discovered that an oxide film of Ta2O5  prevented the flow of current when tantalum was connected as an anode in solutions. This film  however  makes tantalum an excellent electrolytic valve for use in the rectification of alternating current  and the first large commercial application of the metal was in the Balkite battery charger  used extensively in the radio receiving sets of the 1920 s.

Occurrence and Sources

Tantalum ranks fifty fourth in order of concentration of elements in the earth s crust  and is definitely an uncommon metal. It is always found associated with columbium  which is about eleven times as prevalent. The most important mineral source is a ferrous manganese tantalate columbate  (Fe Mn)(Ta Cb)2O6. If the Ta2O5 content exceeds the Cb2O5 content the mineral is called tantalite  if the reverse is true it is called columbite. These minerals are usually found in pegmatite dikes in quantities which seldom exceed a few pounds per ton.

Tantalum is also present in other minerals  such as pyrochlore  fergusonite  samarskite  euxenite  and polyerase. In most of these the columbium content exceeds that of tantalum. As the demand for columbium grows  the processing of high columbium low tantalum ores should make available increasing amounts of tantalum concentrates from these sources. In fact  a good deal of the tantalum has been and continues to be derived from concentrates higher in columbium than tantalum because of the relative scarcity of tantalum concentrates.

Tantalite has been produced chiefly in western Australia where it was mined from alluvial deposits in the Pilbarra District (now a minor source)  in eastern Brazil  Nigeria  and the Belgian Congo Tantalum concentrates are being produced as by products of tin (cassiterite) placer mining operations in the Belgian Congo and to a minor extent in Malaya. Currently  concentrates of this type from the Belgian Congo account for one third of the United States imports of tantalite concentrates.

The United States has only minor and scattered deposits of tantalite  tantalum ore mining is almost entirely a foreign industry. More than 99 per cent of the United States tantalite supply is imported. In 1958 some 84 per cent of the imports came from the Eastern Hemisphere. This dependence on water borne ore imports is one factor opposing the greater use of tantalum.

The ores are concentrated by hand separation  washing  tabling  and electrostatic and electromagnetic means. The concentrates as received ordinarily contain 60 per cent or more of combined oxides (Ta2O5 and Cb2O5)  and associated impurities are iron  tin  titanium  zirconium  silica  and manganese. In the concentrates imported into this country the ratio of Ta2O5 to Cb2O5 varies from 12 1 to 3 4  with the average probably about 1 1 for the bulk of the material.

Production and Price Statistics

Data on the domestic shipments of columbium tantalum concentrates  imports of concentrates  ore consumption. It will be noted that the quantities of metal produced are much less than the quantities contained in the ores consumed. This is because the greater portion of both columbium and tantalum concentrates is fed directly to electric furnaces to produce ferrocolumbium and ferrotantalum columbium. These alloys are used in some austenitic stainless steels where about one per cent columbium and tantalum acts to prevent carbide precipitation.

While prices for imported tantalite are not quoted publicly  material containing approximately 60 per cent. Ta2O5 sold in 1959 for 83.50 4.25 per pound of contained tantalum pentoxide. This represented a decline from the prices in the neighbourhood of 86.25/lb at the beginning of 1958.

Prices for tantalum metal have boon decreasing in recent years as the number of producers and the total plant production capacity have increased. As of late 1960 tantalum powder was generally available at about $40 per pound  welting stock at $35 per pound  and sheet at $55 60 per pound.


In addition to the usual problems encountered in separating a desired metal from the other components of a given raw material  the extraction of tantalum from ores or concentrates is even more difficult because of the presence of its chemically similar sister metal columbium. In this respect tantalum and columbium are quite like zirconium and hafnium insofar as the problem of complete separation of the two is concerned.

Fractional Crystallization. The classical separation method of Marignac  in use since the early 1920 s  has been almost  completely superseded by solvent extraction methods in recent years. The fractional crystallization process as practiced industrially consists of these steps

Pulverized (about 200 mesh) tantalum concentrates are fused with sodium hydroxide in a continuous furnace to form crude sodium tantalites and columbates.

Cooled flakes of the fusion mass are leached with hot water and then with hydrochloric acid to remove most of the iron  manganese  silica  tin  and titanium impurities. Insoluble tantalic and columbic acids (hydrated oxides) are formed by this treatment.

The mixed tantalic and columbic acids are dissolved in hydrofluoric acid  and sufficient potassium hydroxide or fluoride or carbonate is added to form a solution of potassium fluo tantalate  K2TaF7  and potassium columbium oxyfluoride  K2CbOF5.

After the hot solution has been filtered to remove insoluble matter the filtrate is allowed to cool in a crystallizer. Potassium fluotantalate  which has a solubility of 7.5 gpl in water at room temperature  precipitates from the solution  while the K2CbOF5.H2O  which has a solubility of 91.5 gpl in water  remains in solution.

The K2TaF7 crystals are separated from the slurry by filtration and dried in a steam heated tray dryer.

Tantalum powder is obtained by electrolysis of fused K2TaF7 or by sodium reduction of this salt.

The above steps are primarily batch operations  and the fractional crystallization process is better suited to production of pure tantalum than of pure columbium. It is for these and other reasons that this process has been largely abandoned as a commercial process in recent years in favour of the liquid liquid extraction method of extracting and separating tantalum and columbium from ores.

Liquid Extraction. The present widely used method of separating tantalum and columbium by liquid extraction grew out of work conducted in the early 1950 s by several investigators  chiefly those of the U.S. Bureau of Mines  Albany  Oregon and the Ames Laboratory of the U.S. Atomic Energy Commission. The basic principles relate to the effect of acid concentration on the relative solubilities of tantalum  columbium  and other metal fluorides in aqueous and methyl isobutyl ketone systems.

When an acidic aqueous solution of the fluorides of tantalum  columbium  and the other metals present in the mineral concentrates is placed in contact with methyl isobutyl ketone  the tantalum fluoride is extracted by the organic phase at a low acidity and the columbium fluoride at a high acidity. By contrast  the other fluorides tend to remain in the aqueous phase. The solubilities are so different that clean extractions of very pure tantalum and very pure columbium can be obtained in a relatively few extraction stages.

In industrial practice  the fine ground ore (–200 mesh) is digested with concentrated hydrofluoric acid  prepared from anhydrous HF and deionized water  in tanks lined with polyethylene or Haveg. The tantalum and columbium oxides dissolve while the bulk of the impurities remains in the gangue. The tantalum columbium liquor is separated from the undissolved solids by filtration and/or decantation and pumped to a holding tank. From here on the processes differ somewhat.

In one plant a stepwise extraction is used. The tantalum columbium feed solution is adjusted to a low acidity and fed to the polyethylene extraction cascade where it is treated with methyl isobutyl ketone in mixers and settlers operating in tandem. The tantalum is first extracted into the organic phase at the low acidity and then stripped from that phase by deionized water to yield an aqueous solution of extremely pure tantalum fluoride. The tantalum free feed solution is acidified to a high acidity with sulfuric or hydrochloric acid and treated with pure methyl isobutyl ketone in the columbium extraction section of the cascade. Here the columbium content is extracted into the organic phase  leaving essentially all the remaining dissolved impurities in the aqueous phase. The organic phase is then treated with deionized water to remove columbium fluoride from the ketone as a very pure material in aqueous solution.

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