Proceedings of the International Symposiumon

URANIUM TECHNOLOGY

VOLUME I

BHABHA ATOMIC RESEARCH CENTRE,TROMBAY, BOMBAY, 400 085

DECEMBER 13-IS, 1989

Organised byENGINEERING SCIENCES COMMITTEE

BOARD Of RESEARCH IN NUCLEAR SCIENCESDEPARTMENT OF ATOMIC ENERGY

GOVERNMENT OF INDIA

SYMPOSIUM ORGANISING COMMITTEE

1. Shri S. Sen, BARC - Chairman

2. Shri R.K. Garg - IRE Ltd, Bombay

Z. Shri M.K. 8atra - UCIL Jaduguda

4. Shri J.L. Bhasin - UCIL Jaduguda

5. Shri K. Balaramamoorthy, - NFC Hyderabad

6. Shri P.R. Roy - BARC

7. Shri A.N. Prasad - BARC

8. Dr.R.M. Iyer - BARC

9. Shri T.K.S. Murthy - IRE Ltd.

10. Prof. S.L. Narayanamurthy - IIT, Bombay

11. Shri T.A. Menon - FACT, Cochin

12. Comdo. K.C. Chatterjee - DCL Bombay

13. Shri C M . Das - NPC, Bombay

14. Dr.C.K. Gupta - BARC

15. Shri K.S. Koppiker - BARC

16. Dr. Ashok Mohan - BARC

17. Dr.V. Venkat Raj - BARC

18. Shri G.R. Balasubramanian - JCCAP.

l<*. l»r.G. Viowan.3th.in - HMD.Hyderabad

20. Chi i M.R. Balakriahnan fennc

I ' l . Uhri u.R. Marw.ih - Member .Secrotary

TECHNICAL COMMITTEE

1. Shri K.S. Koppiker, 8ARC - Chairman

2. Shri V.S. Keni, BARC

3. Shri S.K. Chandra, IRE Ltd, Bombay

4. Dr.T.K. Mukherjee, BARC

5. Or.A. Ramanujam, BARC

t>. Shri S.N. Bagchi. BARC

7. Shri U.ft. Marwah, BARC - Member Secretary

The National Symposium on Uranium Technology was held atBARC, Bombay during Dec. 13-15,1989 under the auspices of Engi-neering Science Committee of Board of Research in NuclearScience, Department, of Atomic. Energy.

In the context of expanding nuclear power programme inIndiai the need for production of large quantity of uranium fuelfrom indigeneous resources hardly needs any elaboration. Afterthree decades of experience in this area, it is thus appropriateto pool together the expertise and experience gained in thecountry during this period in various aspects of uranium technol-ogy like exploration, mining, ore processing, refining and con-version to oxide and metal. This symposium has provided anopportunity to the uranium technologists to interact and exchangetheir experience and plan future strategies.

Being the first symposium on this toppic the response hasbeen excellent as evidenced by the extensive participation andcontribution of papers covering almost all aspects of uraniumtechnology. Altogether 69 papers, including invited lectureshave been presented and the proceeding* have been brought out intwo volumes.

It is our hope that this technical coverage of the pro-ceedings and panel discussion would serve as a valuable referencematerial for uranium technologists in the coming years; We wishto record our sincere thanks to the members of the organising andvarious other committees, authors of invited lectures, andcontributed papers, and panel members for making it possible tobring out this proceedings. The cooperation extended by Head,Library and Information Division, BARC is gratefuly acknowledged.

TECHNICAL COMMITEE

CONTENTS

INAUGURAL SESSION

1. Introductory Remarks by Al

Shrl S. Sen, Chariman, Organizing Committee

2. Welcome Address by AS

Dr. P.K. Iyengar, Director, BARC

3. Presidential Address by A7

Dr. M.R. Srlnivasan, Chairman, AEC

4. Inaugural Address by A13

Dr. H.M. Sethna, Chairman, TOMCO 4 Tata Electric Companies

5. Vote of thanks by A18

U.R. Marwah, Member Secretary, Organizing Committee

6. Keynote Address by A20

Shri R.K. Carg, CMD, IRE Ltd.

TECHNICAL SESSION I

Plenary Lectures

Uranium mining In India - A38

Past, present and the future

H.K. Betra, Adviser, UCIL, Jaduguda.

TECHNICAL SESSION II

II A Uranium Prospecting

Contributed Papers 1

Structure as a guide for uranium exploration

In the Turamdlh-Mohuldlh area, Slngbhum Dt.Bihar.

R. Mohanty, M.B. Vcrma

Prospecting for uraniua in carbonate rocks of 19

Vempalli formation, Cuddappah

Basin, Andhra Pradesh

H. Vasudeva Rao, J.C. Nagabhushana, A.V. Jeygopal

and M. Thiaaiah

Evaluation of favourable structural features for 36

uranium f roa airborne geophysical surveys

over parts of Hadhya Pradesh

X.L. Tiku, S.V. Krishna Rao and Bipan Behari

Integrated geophysical investigation for uraniua 49

A case study froa Jaaini, Nest Xaaerg Dt.

R. Srinivas, J.K. Dash, S. Sethuraa, K.L. Tlku, Bipan Behair

K.L. Tlku, Bipan Beharlt

Natural theraoluainescence of whole- 74

rock as a potential tool in the exploration for

sandstone type uranium deposits: Application to

the lower Mahadek sandstone of Meghalaya

R. Dhana Raju, R.C. Bhargava, A.Paneerselvaa

and S.M. Vlrnave

Hydrogcochealcal exploration for uraniua: 90

A cast study froa the Cuddappah Basin, Andhra Pradesh

R.P. Singh, P.X. Jala, B.l.H. Kuaar, S.S. Rao,

A.V. Patwardhan and S.C. Vasudeva

An Alpha-gaaaa counting integrating device 109

for uraniua exploration

G. Jha, M. Gaghavayya, H.H. Srlnlvasan, S. Sastry.

Ceostatlstlcal study of Bhaten ore deposit

C.V.L. Bajpal and P.P. Shai

II B Analytical Techniques in Uranium Technology

Uranium analysis using an on-lone background 147

correction programme with carrier-distillation

technique by a computer controlled spectrometer

R.K. Dhumwad, A.B. Patwardhan, V.T. Kulkarni,

K.Radhakrislinan

Deterainatlon of trace metals in uranium oxide 152

by ICP-MS

S. Vijayalakshmi, R. Krishna Prabhu, T.R. Mahalingam

and C.K. Mathews

Development of flow-injection analysis technique for 157

uranium estimation

A.H. Paranjape, S.S. Pandit, S.S. Shinde,

A. Ramanujam and R.K. Dhumwad

Standardisation of DC Arc carrir-distlllation 166

procedure on a direct reading spectrometer

for the deterainatlon of B, Cd etc. in nuclear grade uranium

S.S. Biswas, P.S. Murty, S.H. Msrathe, A. Sethuaadhavan

V.S. Oixit 1. Kalaal and A.V. Sankaran

Spectrographic determination of B,Cd and Nl in 182

aagnesiua fluoride

A. Sethuaadhavan, V.S. Dixlt and P.S. Murty.

Estiaatloo of uranlua in leach liquors of low 189

iron content: Modification of a spectro-

photoaetrlc method using *-(2-pyridil a 20) resorclnol

G. Suryaprabhavati, Leela Copal, G.S. Chawdary

and Radha R. Das.

Discussions 200

TECHNICAL SESSION III

III A Mining and Ore Benefeclation

Contributed Papers 204

Development of mining at Jaduguda

J.L. Bhasln

Role of support services at Jaduguda uranium mine 232

S.D. Khanwalkar, V.N. Radhakrlshnana, M.N. Srinivasan,

Pinaki Roy, S.N. Bannerji

Recovery of uranium concentrate from copper tailings 254

S. Chakraborty, U.K. Tiwari and K.K. Beri

Significance of petrology in the ore processing 284

technology with special reference to the uranium

processing from the copper tailings of Singbhua Thrust Belt

N.P. Subrshmanyam, T.S. Sunilkuaar,

D. Naraslahan and N.K. Rao

Improved gravity flow sheet for the recovery of 300

uranium values from the copper tailings

R. Natarajan, R.S. Jha, U. Sridhar, N.K. Rao

Magnetic separation for precoocentration of uranium 318

values from copper plant tailings

R.S. Jha, T. Srinivasan, R. Katarajan, U. Srldhar

and N.K. Rao

Preliminary beneficiatlon studies on uranium ore 332

from Tummalapalli, Andhra Pradesh

N.P.H. Padmanabhan, U.Sridhar, N.K. Rao

Discussions

TECHNICAL SESSION III

111 H An.Tlytic.il Tochniques in Uranium Technology-II

Contributed Papers 349

Rapid determination of uranium in uranyl

nitrate solution by Ganma Spectronetry

T.K. Shankaranarayanan and D.S. Gupta

Modification of fluoriaetric Method of uranlua 356

analysis for Jaduguda Plant Saaples

A.B. Chakraborty and V.H. Pandey

Determination of uraniua In sea water by 369

adsorptlve differential pulse voltaaetry

R.N. Xhandekar and Radha Raghunath

Difficulties In preparing a standard saaple 376

of uraniua aetal having traces of nitrogen

R.S.D. Toteja, B.L. Jangida, N. Sundaresan

Estlaation of Bangancse in tailings plant 382

effluents by ICP-AES

Joydeb Roy and V.H. Pandey

Voltasaetric studies of uraniua (VI) 389

reduction

Discussions 402

TECHNICAL SESSION IV

Uranium Ore Process Technology

Invited Lecture

Technologies for processing low-grade uranium 403

ores and their relevance to Indian Situation

T.K.S. Murthy

Contributed Papers

Jaduguda Uranium Mill-Rich experience 431

for future challenges

K.K. Berl

Grinding and leaching characteristics 463

of the Indian uraniua ores.

V.M. Pandey and R.U. Choudhary

Recovery of uraniua by direct low-acid 477

leaching froa copper concentrator tailings

V.M. Pandey, R.U. Choudhary, A.K. Sarkar,

A.P. Bannerjee, A.B. Chakraborty, N. Malty

Selection of ion exchange resin for uraniua 485

adsorption froa Jaduguda leach liquors

D.P. Saha and V.M. Pandey

Apprlicatlon of advanced technologies for uraniua 498

•inlng and processing at Narwa Pahar and Turaadih Projects

R.C. Purl and R.P. Veraa

Impounding of tailings at Jaduguda - 528

Planning, design and aanageaent of tailings daa

S.N. Prasad and K.K. Beri

TECHNICAL SESSION V

Uranium Ore Process Technology-contd

and Byproduct Uranium

Contributed Papers

Nuclear pure uranium from ores using weak 555

base ion-exchange resins

S.V. Parab, S.S. Charat, G. Cherian and K.S. Koppiker

Development of an Integrated process for recovery 570

of uranium from ore and its refining at the

location of new uranium mill at Turamdlh

R.A. Nagle, S.V. Parab, S.S. Charat, A.B. Giriyalkar

and K.S. Koppiker

Preparation of nuclear grade uranium oxide 582

from Jaduguda leach liquor

V.M. Pandey, A.B. Charkraborty and N. Haity

Uranium recovery from phosphoric add 592

G. Sivaprakaah

On-site teats for recovery of uranium from wet 621

process phosphoric acid at FACT

H. Singh, R.A. Hagle, A.B. Giriyalkar, M.F. Fonseca

and K.S. Koppiker -

Recovery of uranium from nltro-phos acid 628

R.A. Nagle, A.B. Giriyalkar and K.S. Koppiker

Recovery of uranium from monazlte - a fresh 635

look at the current practice

S.L. Mishra and K.S. Koppiker

Recovery of uranium from sea water- 643

A laboratory study

D.V. Jaywant, N.S. Iyer and K.S. Kopplker

TECHNICAL SESSION VI

Uranlua Refining

Contributed Papers

Operating experience In the refining of uranium by 653

solvent extraction using sixer settlers

SMT. S.B. Roy, H. Singh, K. Kuaar, A.M. Meghal,

V.N. Krishnan, K.S. Koppiker

CALMIX-Innovatlve aixer-settler systea 659

C.K.R. Kaiaal, B.V. Shah, I.A. Siddlqui, S.V. Kuur

Precipitation of aaaonlua dluranate-a study 666

T.S. Krishnaaoorthy, N. Kahadevan, H. Sankar Das

Continuous reactor systea for precipitation of 688

uraniua froa uranyl solution

I.A. Siddlqui, B.V. Shah, S.H. Tadphale, S.V. Kuaar

Preparation of aetal grade uraniua trloxide 695

through aaaoniua diuranate precipitation route

S.R. Raaachandran, P.D. Shrlngarpure and A.M. Meghal

Studies on preparation and characterisation of 701

aamoniua uranyl carbonate (AUC)

V.N. Krishnan, M.S. Visweswariah,. P.D. Shringarpure

and K.S. Koppiker

Batch precipitation technique-a process for 708

U(>2 powder procutlon.

A.K. Srldharan, G.V.S.R.K. Somayaji, N. Swaminathan

and K. Balaraoamoorthy

Development of AUC route for production of U0_ Powder 712

U.C. Gupta, Smt. Meena R. and N. Swaminathan

Analytical technique in uraniua dioxide 728

fuel production stream.

T.S. Krlshnan, S. Syaasundar, B. Gopalan,

R. Narayanaswaay and C.K. Raaaaurthy

TECHNICAL SESSION VII

Uranlua Metal Production

Contributed Papers

Iaproveaents in process technology for 750

uraniua aetal production at UMP

A.H. Meghal, H. Singh, A.V. Vedak, K.S. Koppiker

laproveaents In equipaent design for 756

hydrofluorinatlon of UOjto UF^

A.V. Vedak, R.N. JCerkar, and A.M. Meghal

Hagnesio-theralc reduction of Ufy to 762

uraniua aetal - plant operating experience

S.V. Mayekar, H. Singh, A.M. Meghal,

K.S. Koppiker

Recovery of uraniua froa aagnesiua fluoride 770

slag at UMP

P.K. Bandopadhyay, B.M. Shadakshari,

H. Singh and A.M. Meghal

Future trends In the processing of 777

uraniua slag generated suring production of uraniua

aetal.

Keshav Chandra, Mahesh Singh, II. Singh, A.M. Meghal,

K.S. Koppiker and S. Sen <*

Quality assurance during uranium metal production at 790

UMP

V.N. Krishnan, R.D. Shukla, M.S. Visweswariah

Novel surface chemical treatment to improve the 796

quality of scintered U02 pellet

B. Venkataramani and R.M. Iyer

P.C. based uranium enrichment analyser 805

V.K. Madan, K.R. Gopalakrishnan and B.R. Bairi

Discussions. 810

TECHNICAL SESSION VIII

Environmental aspects, Health fc Safety

Contributed Papers

Treatment of uraniua tailings vis-a-vis radius 811

containment

P.M. Markose, K.P. Eappen, H. Raghavayya, K.C. Plllal

Radon problems in uraniua industry 833

A.H. Khan and M. Raghavayya

Effective dose evaluation of uraniua mill workers at 848

Jaduguda

G. Jha and M. kaghavayya

Radiological and envlronaental safety aspect* of

uraniua fuel fabrication plants at Nuclear Fuel

Coaplex at Hyderabad

S. Viswanathan, B. Surya Rao, A.R. Laxaan and T.

Krishna Rao ^

Litaits of plutonium contamination in reprocessed 865

uranium for handling in natural uranium plants

V.K. Sundaram and M.R. Iyer

Biosorption of uranium by yeast 874

A.K. Mathur, N. Huralikrishna, V. Krishnaaurthy and

R. Sankaran

»

Discussions 885

TECHNICAL SESSION IX

Health and Safety Aspects-contd

General Cheaistry of uranlua technology

Contributed Papers

Operational health physics experience at uranlua 888

aetal plant, Troabay

P.P.V.J. Naabiar, Pushparaja, J.V. Abrahaa

Radio activity levels in the process streaas of 897

uranlua aetal plant (UHP) at Troabay

Pusbparaja, S.G. Sahasrabhude, J.V. Abrahaa and M.R.

Iyer

Radiological and conventional safety aspects of 902

aachlnlng operations of uranlua Ingots

V.B. Joshi, I.K. Ooaen, S. Sengupta, T.S. Iyengar

Radiation risks, aedlcal survsillaacc prograaae and 910

radiation protection in the alning and ailllsg of

uranlua ores

Dr. A.K. Rakshit

Separation of uranlua VI, Chroalua and zlrconlwB by 924

solvent extraction with crown ethers

N.V. Deorkar and S.M. Xhopkar

Uranyl ion transport across tri-n-butyl phosphate-n -39

dodecane liquid aembranes

J.P, Shlkla and S.K. Mlshra

TECHNICAL SESSION X

Project Management

Znvited Lecture

Consultancy, project engineering service for the 947

uraniua industry

A.K. Bhattacharya, Vice Chairaan DCL

Contributed Papers

Project Manageaent-progleas in execution 958

D.C. Nalr, FACT

Conditions required for opening of a coaaerclal 985

•lneral deposit

S. Sastry, UCIL

Management of uraniua aining and process wastes at 998

Turaadih Project

R.C. Purl and R.P. V e m , UCIL

TECHNICAL SESSION XI

Panel discussion on

/

"Present status and future strategies on uraniua

technology"

Al

INTRODUCTORY REMARKS BY S. SEN

CHAIRMAN, SYMPOSIUM ORGANIZING COMMITTEE

Dr. H.N. Sethna, former Chairman, Atomic Energy Commission and Principal

Secretary to the Department of Atomic Energy, Dr. Srinlvasan, Chairman,

Atomic Energy Commission, Dr. Iyengar, Director, Bhabha Atomic Research

Centre, Shri Garg, Chairman and Managing Director of the Indian Rare Earths

Limited, Shrl Marwah, Secretary of the symposium organising committee, my

dear colleagues, distinguished delegates, ladies and gentlemen,

On behalf of the Symposium Organising Committee, it gives me great pleasure

to extend a very hearty welcome to you all on the occasion of the

inauguration of the symposium on "Uranium Technology".

When the Board of Reseerch in Nuclear Sciences of the Department of Atomic

Energy wanted from me suggestions on subjects for symposium, the topic of

uranium technology came up because of three reasons. Firstly, the year

1989 marks the bicentenary of the discovery of uranium. The second reason

was my long association with uranlua technology, first in the Uranium Metal

Plant during 1956-63, then at the Uranium Hill at Jaduguda during 1964-70

and finally mt BARC from 1971 onwards. Thirdly, the government has

embarked on an ambitious expansion of the nuclear power programme to 10,000

MWe generation capacity by the year 2000 A.D. A ten-fold expansion of

uranium mining, milling and refining will be required to meet the demand on

fuel material. It was, therefore, felt that we should have a symposium on

"Uranium Technology" at this juncture. I am happy to say that the BRNS

readily agreed to the holding of this symposium under its auspices when we

proposed the topic to them. BARC was chosen as the venue being the birth

place for sost of the uranium production processes.

The element uranium was discovered by the German Chemist Klaprolh in 1769

and was named to commemorate the planet uranus which had just then been

discovered. It was of little commercial Importance till the advent of the

atomic age. Uranium today is the primary fuel In the nuclear reactors and

so far as India is concerned for the first stage of our nuclear fuel cycle

strategy. Uranium is recovered from ore by hydrometallurgical processes

involving acid leaching, ion exchange or solvent extraction and finally

precipitation as "yellow cake". Refining of uranium to nuclear purity is

achieved by solvent extraction using TBP. The annual world production of

uranium concentrate is estimated to be around 40,000 tonnes of uranium

oxide. As far as uraniua fuel is concerned, India is self-reliant today.

Uranium technology is also a trend setter for the development of several

techniques utilised In metallergical and chemical engineering practice, for

example, heap leaching, bacterial leaching, solvent extraction,

ion-exchange, waste management, pollution control etc. The spin offs from

this technology has revolutionised the metal extraction for a large number

of metals like copper, cobalt, rare earths, platinum group metals etc.

Although these advances have been incorporated in practice abroad, they are

yet to be introduced in India.

A review of the uraniua exploration and mining scenario indicates the

urgency for stepping up the programme of uraniua exploration and taking

steps to open new aines at an accelerated pace. Accelerating the programme

for exploration and Mining could result In identifying additional and

perhaps richer uraniua resources. It is necessary that geologists, mining

engineers and cheaical engineers have knowledge of a large nuaber of

processes and equipment for studying a concrete case and optimising all the

conditions of developaent of deposit. It aay be noted that each ore body

constitutes a separate case by reason of geological paraaeters inherent to

the location of the deposit, the physical and chealcal nature of the

gauge, the reserves It represents and the ore grade of the deposit. It is

necessary to adopt an unbiased approach to the study while at the saac tlac

taking as basis the aost well-tried industrial experiences available

elsewhere.

Our uraniua ore grades are low and resources are Halted. Therefore, we

have to make all our efforts to recover uraniua froa all available sources

from copper tailings, from phosphoric acid, from aonazite and perhaps even

from sea water.

A3

If we look at the global picture in respect of uranium technology rapid

changes have taken place in the last two decades in process and equipment

used for uranium production. Many new methods are under study on a

laboratory or pilot plant scale which may altar present practices

altogether. Mention may be made of some of the recent developments

elsewhere in the world, namely, thick puJp leaching including concentrated

acid leaching, high temperature and higher concentration alkaline leaching,

use of horizontal belt leaching and filtration, resin-in-pulp extraction,

fluidized bed precipitation, moving bed and fluidized bed reduction and

hydrofluorination, drying by atomisation, the AUC process, the Excer

process, the Fluorox process, continuous metal production, direct reduction

of UFj, to U0z etc. This is, therefore, the right time for uranium

technologiests to update information, to review experiences on existing

process and equipment and make decisions on modifying the processes,

upgrading the equipment or altogether changing the processes or equipment.

Thus the symposium is being held at an appropriate stage.

The symposium programme Includes topics such as uranium prospecting,

mining, ore benefielations ore processing, refining, metal production,

analytical techniques, health, safety and environmental aspects and project

management. There are one keynote address, seven invited lectures and

seventy four contributed papers. It Is requested that those presenting the

papers may kindly cover the presentation within the time allotted so that

sufficient time is available for discussion. The panel discussion on the

last day will be on "Present Status and Future Statagies on Uranium". It

is hoped that information presented and discussions held in this symposium

will be helpful towards achieving our goals. Being the first symposium on

this subject, it has not been possible to include in detail many of the

topics related to uranium technology. It is proposed to cover these topics

in detail in subsequent seminars. I must apologise for any short-comings

in arranging for accommodation and transport to the participants.

A4

We are deeply grateful to Dr. Sethna for being with us this morning. When

the question of Inauguration of this symposium came up before the symposium

committee the choice was very obvious. We could not think of any other

person except Dr. Sethna to inaugurate this symposium in view of the fact

that the development of all process and design of uranium production plants

in BARC and in other Units of DAE right from the begining were carried out

under his personal guidance. We were also sure that in view of his deep

interest in this subject he would agree to our request. We are indeed

thankful to hi* for sparing his valuable time for this inaugural function.

We are happy to have Dr. Srinivasan who readily agreed to preside over

this inaugural session. We are thankful to Dr. P.K. Iyengar, Director,

BARC who cancelled an outstanding engageaent elsewhere in order to be with

us today. I as also thankful to Shri Garg for agreeing to deliver a

Keynote Address. We are happy that a number of distinguished scientists

and engineers engaged in various aspects of uranium technology are

participating in this symposium) and some have agreed to give invited talks.

1 am happy to note that some persons from the academic institutions are

also attending this symposium. I can also see a number of old colleagues

present in this symposium to share with us their valuable experience. I

take this opportunity to thank all of you including the speakers and the

sessions chairmen and of course my colleagues- in the Organizing Committee,

the Technical Committee and Local Hospitality Committee particularly

Shri Kopplker, Chairoan of the Technical Comaittec whose untiring efforts

made what this symposium is today.

With these words and with genuine hope that we arc going to have a fruitful

syaposium, I would request Dr. Iyengar, Director, BARC to address the

gathering.

A5

WELCOME ADDRESS

BY

DR. P.K. IYENGAR, DIRECTOR, BARC

Dr. Srinivasan, Dr. Sethna, Mr. Sen, Mr. Garg, participants,

distinguished guests, ladies and gentlemen,

It is indeed oy pleasant duty this morning to welcome you all to this

symposium on Uranium Technology. Mr. Sen pointed out that this is the

flrsc time uranium technology is being discussed in a symposium of this

magnitude. The sain reason is, of course, that it is only the Department

of Atomic Energy which is interested in producing large quantities of

uraniua. Uranium technology involves many of the new techniques

especially in fluoride chemistry and fluorine chemistry which really

evolved as a result of research and development in uranium technology.

However, in India uranium has got to be processed sooner or later from

very weak sources like from sea water and ores of very low grade.

Besides, uranium has to be recovered from irradiated fuel. The result is

that we have a complex problem of extracting uranium from very low grade

ores as well as from processed fuel. It is, therefore, appropriate that

this symposium discusses all aspects of the technology including the

economics of each process and the relative merit of process compared to

the other. It is fortunate that we have with us Dr. Sethna who

originated this technology in this country in the Department of Atomic

Energy. Uraniua at one time was considered good for nothing other than

as ballast in ships. But the advant of atomic energy made it a very

important material, and proficiency in uraniua technology became an

loportant factor in the assessment of technological capabilities of

various countries. Fortunately for us, through the initiative of Dr.

Sethna we have mastered all aspects of uraniua technology and of

recovering its by-products to a level in which we could be proud of. We

can confidently plan for the expansion of uraniua technology to aeet all

requirements of the 10,000 MWe nuclear power prograaae in the country. I

remember some of the early days In which this work win being done under

A6

the direction of Dr. Sethna, and I distinctly remember that one of the

characteristics of involving oneself in this new technology, which was

not easily accessible was to have a dare devil psychology in addition to

doing good technological development. It was necessary, and through Dr.

Sethna it was possible to appreciate and encourage this aspect of

evolving a new technology in this Centre. I remember the days when we

worked with fear of a small explosion in a laboratory, which finally

ended up in producing an ingot of uranium, which had the shape of a

Shivalinga and had the power of Shiva as both in energy production and in

'destructive aspects. I am glad that Dr. Sethna is with us today to give

the key-note address on this occasion. No doubt this is an area of

research which is continuously being revived because of economic

considerations and due to the fact that It Is becoming more and more a

strategic material from the point of view of the economic development of

any country. Therefore It Is all the more important that we must have

cooperation and consolidation of our previous efforts in this new area.

I congratulate the BRNS for having organized this symposium at an

appropriate time when methodologies are being evolved and perhaps it will

enable us to achieve a much faster growth of nuclear energy.

Thank you very much for your attention.

A7

PRESIDENTIAL ADDRESS BY DR. M.R. SRINIVASAN

CHAIRMAN, ATOMIC ENERGY COMMISSION &

SECRETARY, DEPARTMENT OF ATOMIC ENERGY

Dr. SeLhna, Shri Sen, Shri Garg, participants to the Symposium, Ladies and

Gentlemen,

I would like to take the opportunity today of discussing some issues of

nuclear power which have received attention of the media both here and

abroad. These concern reports that the United Kingdom has essentially

decided not to go ahead with its Pressurised Water Reactor programme. As

many of you know, the U.K. decided to proceed with the construction of a

PWR of 1175 MWe capacity, named as Sizewell 'B*. This reactor was to be a

prototype of the PUR line and the Central Electricity Generating Board was

in the process of obtaining clearances for constructing additional units at

Hinkley Point.

I was in Vienna a few weeks ago to attend a Senior Experts Group meeting

convened by the Director General, International Atomic Energy Agency. One

of the members of this Group was Lord Walter Marshall who was until

recently the Chairman of the Central Electricity Generating Board, U.K. and

was slated to take over as Chairman of the National Power Company. His

presence at the Senior Experts Group meeting afforded me and other members

of the Croup, an opportunity to get a first hand account of the

circumstances that led to the decision In the U.K. of not proceeding with

additional PWRs and as a consequence, the resignation of Lord Marshall.

The Thatcher Government has had privatisation as an Important part of its

political platform. As a part of this policy telephone services and gas

supply which were earlier state owned monopolies. There has been criticism

amongst an influential section of the Conservative Party politicians that

replacement of a publicly owned monopoly by a privately held monopoly was

not adequate and that it was essential to introduce competition in the

provision of services such as telephones, gas supply, electricity and even

A8

water supply. Bearing this criticism in mind, the framework on

privatisation of the electricity industry brought about an unusual

situation whereby it was not obligatory for the electric utility to ensure

electric supply to customers. Neither was it feasible for the electric

utility to adopt costing principles that would adequately allow returns on

long term investments which characterise nuclear power development. In

simpler terms, during the days when the Central Electricity Generating

Board operated as a public utility it had the territorial franchise for

supply of electricity in England and Wales and it was a monopoly. This

situation is not unusual with electric supply utilities around the world.

They have grown as monopolies in the public sector or in the private

sector. Examples of monopolies in the public sector are Electricite de

Prance, Ontario Hydro, Quebec Hydro etc. Monopolies in the private sector

which have equally successfully fulfilled supply obligations to their

customers are ToJcyo Electric Company and a number of other Japanese

utilities.

Lord Marshall had warned the British Government about complexities that

would be introduced in privatisation of the electric supply industry and

•ore especially about the consequences of removing the monopoly position

that the electric supply industry enjoyed. His objection was not

against privatisation per se. In fact he stated categorically that the

electric supply industry, as a fully private Industry, could still plan

future generation programme in a rational manner, taking into account

all alternative sources of generation, when it continued as a monopoly.

I would like to briefly refer to another aspect of the U.K. programme,

namely, the presently perceived highly uneconomic operation of the

Magnox reactors (Carbon dioxide cooled graphite moderated natural

uranium fuelled reactors). These reactors which formed the first part

of the U.K. programme have indeed been looked upon as a work horse of

the U.K. electric supply Industry for a couple of decades. In fact in

the past they produced and sold electricity much cheaper than from coal.

They also played a very important role in maintaining supply of

electricity during the long coal miners strike. However, in recent

A9

times, the economics of these reactors has suddenly turned unattractive.

The reason for this is related to the presently assessed high cost of

reprocessing of spent fuel. The fuel used in Magnox reactors has

relatively low burn-up, namely 3000-4000 MWe per day/tonne; compared to

6500-8000 MWe per day/tonne for heavy water reactors and about 30,000 to

35,000 MWe per day/tonne for light water reactors. In other words, for

the same quantity of electricity generated, Magnox reactors produced

much larger quantities of irradiated fuel involving much higher

expenditure in reprocessing and waste management. Secondly, the power

density of the Magnox reactor is extremely low. The implication of

this is that at the end of life of the Magnox reactors, a reactor with

250 MWe output leaves behind about 500 tonnes of spent fuel. Compare

this to the incore inventory of a heavy water reactor of equal capacity

which is less than 50 tonnes. The light water reactors have even lower

incore inventories for the same output. Now at the end of life of these

reactors, it is necessary to take the fuel out and reprocess it and

nanage the waste. When appropriate allowances are made for these

activities and the cost of power produced now is loaded for this

purpose, the Magnox reactors become a very expensive proposition.

Another circumstance which entered into the picture is that in the

earlier days of reprocessing of Magnox fuel in the U.K., the plant did

not have adequate waste treatment facilities and there was general

complaint about higher than desirable levels of wastes having been

discharged into the Irish Sea. Some years ago, extensive modifications

were carried out to overcome these weaknesses and these all have added

to increased capital costs for reprocessing and increased operating

costs. In addition, when the privatised National Power Corporation

insists on fixed price contracts for reprocessing, as compared to the

earlier cost plus type of contracts with the reprocessing organisation

(namely, British Nuclear Fuels Limited), BNPL has found it necessary to

build In substantial margins for future costs especially those relating

to long term waste management.

A10

One may ask the question, whether the U.K. experience does not apply to

all nuclear p wer. The answer to this is that the French who have a

line of Pressurised Water Reactors (using low enriched Uranium) have a

long history of running the reactors and also in reprocessing. They

find that the reprocessing and waste management costs do not, in fact,

add an unacceptable burden to the cost of power. So far as heavy water

reactors are concerned, Ontario Hydro which has about 10,000 MWe of

operating nuclear capacity, similarly find that the costs related to

management of spent fuel add only to some 4Z of the cost of unit energy.

When the Chernobyl accident took place many members of the general

public intuitively thought that such an accient could take place in any

nuclear installation. It was not easy for them to appreciate that the

particular* kind of reactor at Chernobyl had certain unique infirmities

specific to that type and design of reactor and that the operating

personnel transgressed many of the specific safety provisions.

Similarly when the media discusses the U.K. situation, an impression may

be created that the circuastances that have rendered the U.K. nuclear

power programme unattractive economically are general in nature and

could apply to other cases also. This is certainly not true. In fact

even now there are examples of Prance, Canada, Japan and South Korea, to

mention only some countries, where nuclear power in significant

quantities is being generated both safely and economically.

When talking about energy options, there is a tendency to generalise

from the experience of one country to another. Often the differing

circuastances prevailing in different countries are Ignored. For

example, people ask the question why India should develop nuclear energy

when the United States has stopped building new nuclear projects.

People do not realise that the United States has a very large reserve of

coal, petroleum and gas on Its territory or that the USA has access to a

very important share of global petroleum resources. Similarly the

question will be asked why India should pursue nuclear energy

development when the United Kingdom has recently found nuclear power to

be uneconomical. They do not see that the U.K. has been rather

fortunate in finding very large oil and gas reserves in Its offshore

areas and also that It hats access to the enormous natural gas reserves

All

in the North Sea coming under the control of Norway. We should look at

examples such as France, Japan and South Korea where non-availability of

alternative energy sources has made these countries turn to nuclear

energy. They have met the technological and economic challenges and

have developed safe and cost effective nuclear power. It is ray belief

that India also has the technological and managerial capability to

achieve what has been achieved in France, Japan and South Korea.

I now turn to the subject of this Symposium, namely, Uranium Technology.

From the announcement sheet, I notice that this is the first Symposium

of its kind aimed at dissemination of information, sharing of experience

and identifying areas of technological development relevant to

production of Uranium. There are a number of groups in the Department

of Atomic Energy, especially at the Bhabha Atomic Research Centre,

Atomic Minerals Division, Uranium Corporation of India Limited and the

Nuclear Fuel Complex which are involved in different facets of Uranium

technology.- The country has embarked on a nuclear power programme with

a target of 10,000 MWe to be achieved by the year 2000. This programme

is depending crucially on locating adequate quantities of Uranium within

the country. It is also important to extract this Uranium and convert

it into fabricated nuclear fuel in the most economical manner. There is

also the very important question of minimising the impact on the

environment of Uranium mining and fuel fabrication. I note that this

Symposium will cover all these and other relevant topics.

We are extremely happy that Dr. Horn! Sethna has found it possible to be

with us this morning. All of you know that he as the Chairman of the

Atomic Energy Commission for over a decade has been involved with many

facets of the uranium work. He was personally involved with the setting

up of the Uranium Metal Plant and with the technological aspects of the

Uranium Extraction Plant at Jaduguda. During his stewardship, the

Nuclear Fuel Complex was planned and established. He has also been

responsible for guiding the expansion activities of the Atomic Minerals

Division. In recent years, he has been heading the Tata Oil Mills

Company Limited, Tata Electric Group of Companies, Tata Consulting

A12

Engineeers and a number of other Tata ventures. We could not have had .1

better person than him to inaugurate this Symposium. I now have i',roat '

pleasure in requesting Or. lloml Sethua to deliver tlie inaugural

and inaugurate the National Symposium on Uranium Technology.

A13

INAUCURAL ADDRESS

BY

DR. H.N. SETHNA, CHAIRMAN,

TOMCO AND TATA ELECTRIC COMPANIES

I am happy to be here with you this morning for the inauguration of the

Symposium on "Uranium Technology". Uranium is the ki.y to the nuclear

fuel cycle and uranium technology is an integral part of this technology.

It Is, therefore, ppropriate that the Board of Research in Nuclear

Sciences of the Department of Atomic Energy has sponsored this symposium

In the bl-centenary year of llic discovery of uranluis. The topic is of

personal interest to me because of my association in its early stages.

Some thirty years ago, when Dr. Bh3bha Initiated the development of

nuclear energy, two decisions were taken; the first was to construct the

CIRUS reactor and, second to work on the production of uranium metal fuel

in the country. In the year 1956, the task of producing uranium metal

was assigned to a group called "Project Firewood". This group completed

the process development, design and layo^w of the plant during 1957. The

layout and working drawings of the plant were approved in November 1957;

civil construction, fabrication and erection of equipment were completed

In about a year. 1 still remember the e:.-Jturnout created when the* first

Ingot of nucleur grade uranium metal was produced on January 30, 1939.

A14

Some persons said that this Bade India the first country in Asiaa,

outside USSR to produce nuclear fuel material. I do not think so,

looking back I think it was China.

We entered the technological phase of extraction of uranium from ore when

BARC set up the uranium Hill at Jaduguda for treating 1000 MT of ore per

day. The task was especially challenging as the ore was low grade. The

process flowsheet was frozen based on the work done in the laboratory,

followed by pilot plant scale studies and the complete design of the

plant was carried out by our engineers. The construction of the Hill was

completed In 1967 and it was handed over to the then newly formed Uranium

Corporation after successful commissioning. Even after 22 years this

uill is running to full capacity and has supplied all the uranium

concentrate for research and the PHW power reactors.

I understand that the Atomic Minerals Division has been successful in

proving uranluu reserves In Meghalaya and in the Cuddapah district in

Andhra Pradesh in which the ore Is reported to be of different type from

the Jaduguda ore and may require a different technology. Once a mineral

deposit is discovered and ore resources arc estimated, many wore

investigations are necesuary to make a dcpoult commercially viable. Data

rcj'ardlns rock characteristics, btrliaviour of the ore hotly, liytlrnloj'ic.'il

conditions, extractabLLity of uranium Crow I tie occ, dlujtoual ot mine

water and waste rock and suitable ;;ltes for mill tailings disposal,

A15

besides easy availability of raw materials, water and electricity, are

required to be collected for assessing the suitability of the deposit for

opening a new sine and setting up a mill. There isttherefore a challenge

for our technologists in this field. Apart from the process technology

the problem of logistics may pose another challenge in a location like

Meghalaya.

The extraction of uranium was only one aspect of uranium technology.

Uranium dioxide was to be the fuel for heavy water based nuclear power

reactors. The development work on ceramic grade uranium dioxide

production was Initiated in BARC as early as 1962. The process know-how

was generated by the Uranium Metal Plant Group. This know-how was

employed in setting up a plant at Nuclear Fuel Complex for the production

of ceramic grade uranium dioxide powder and the plant was commissioned in

1971. BARC also developed the process for converting enriched uranium

hexaflourlde to uranium dioxide based on which a plant was also set up at

the NFC. I understand that at NFC, several Innovations have been made in

the process technology since then. Similarly, I understand that for the

new mill coming up at Turamdih, UCIL has opted for belt filtration,

followed by counter-current Ion exchange using undarlfled leach liqour

and finally elutlng uranium with dilute sulphuric add. Once add

elution Is selected It would be advantageous to go In for ELUEX process

using amine solvent extraction route. This would help In overcoming the

silica waste problem faced during refining. However, we should not feel

A16

satisfied with these achievements. Improvements in equipment design and

process technology have to keep pace with developments in other countries

of the world. In the task of technology up-gradation, some of the first

generation experts who have retired or would retire soon from service

could be utilised.

After reviewing our achievements, this is also an appropriate tine for

appraising reality. We have to live with the fact that our ores are low

grade and resources are Halted. Therefore, development of economic and

efficient processes is imperative. We have also to sake all our efforts

to recover uranium from any available source. One such source is the

recovery of uranium from the wet process phosphoric acid production and

from copper tailings. I understand that the first plant for recovery of

uranium *rom this source is to be set up at FACT, Cochin. If exploited

properly, phosphoric add plants could be a perennial source of uranium.

India's requirements for phosphatlc fertilisers is increasing every year,

and the strategy of buying phosphoric acid from abroad may change in the

near future and more plants may come up to produce the acid in the

country. This would further increase the uranium availability from the

source. As long as the cost does not exceed the cost of production from

a newly developed uranium mine in our country, we should go In for

setting up plants for uranium recovery from copper tailings and from

phosphoric acid, irrespective of their size. Again such decisions have

to take into consideration availability of manpower and financial

constraints.

A17

To sum up, although it has been an eventful journey in the last three

decades, there has to be greater thrust on innovation and timely

completion of projects for uranium production for meeting the increasing

demand for the projected nuclear power programme. Emphasis on R & D has

to be maintained and the young engineers and scientists have to come up

with new ideas because the perspective has changed in these three

decades. Earlier it was self-sufficiency and now it is competitiveness.

The growth in the field in my opinion was the result of undertaking R & D

by our own. If properly carried out such an approach effects more

economics than its costs. 1 hope engineers and scientists in the DAE

would continue this philosophy and bring out better methods of using

India's scarce resources of uranium and meet all the challenges in the

field of uranium technology. To give an example that such an approach

pays is that In the technology for zirconium production we could venture

to take the TBP extraction route in NFC plant although the plants

operating elsewhere in those days had adopted hexone-thiocyanate system.

I hope that this symposium will help consolidating all know-how and

planning out strategies for uranium production in India*

I wish your deliberations all success. I have great pleasure in

inaugurating this symposium on "Uranium Technology".

A18

Vote of Thanks

By

U.R- Marwah, Member Secretary, Organizing Committee

It is a matter of privilege to have been given the opportunity to

propose vote of thanks on behalf of the Organizing Committee.

One of the many things I have not done in my life is to thank

such a galaxy of accomplished people and that too in such an ambience

of the symposium on Uranium Technology. Perhaps it was good that I

did not do it before so that I can thank today, the most genuine

contributors, and thus remain truthful to myself. I have also a

feeling that I am thanking you all on behalf of the nation and

particularly those few who played the sheet anchor role in the

development of Uranium Technology and through that the national

development. Of course, there can be no occassion in the annals of

Atomic Energy without remembering Dr. Homi Ehabha - the visionary -

but it also appears at a time when new decade is to begin the presence

of Dr. Homi Sethna - the doer - has been and shall be dear to us all

who have anything to do with nuclear technology and through that the

national development. Dr. Homi Sethna - we all thank you in agreeing

to inaugurate and grace this symposium.

This occasion when all the illumlnarics of the past and present

generation could be brought together would not have been possible but

for the unstinted support received from all quarters. We thank Dr.

M.R. Srinivasan, Chairman, AEC who In spite of his busy schedule

agreed to preside over this function. We are grateful to Dr. P.K.

Iyengar, Director, BARC who very kindly cancelled his other

appointments to be present here and grace this occasion. I thank Shri.

R.K. Carg, CMD, IRE for agreeing to the request of the organising

committee to give key note address.

I was overwhelmed with the response received from various

sponsors and among them I must* Mention Shri. J.L. Bhasln, CMD, UCIL

and Commodore Chatterjee, DCL, who have been so understanding and

occomnodatlve that it has been a pleasure to interact with them.

A19

Thanks are also due Co Shrl. A.S. Dikshit, HPD, Shri. M.R.

Balakrlshnan, Head, Library and Information Services, PRO's office,

Shri. Subramaniam, A.O., Training School. As the secretary of the

organising committee, it Is ay pleasant duty to acknowledge unreserved

co-operation 1 got from the aeabers of the committee and other

colleagues who worked tirelessly in organising this symposium.

Organizing this symposium, we have tried to do our best but it

may fall short of your expectations because your expectations of our

best may have been high. However, from this moment onwards, the

symposium is ours and not of the organizers. In case of any

inconvenience or organizational problems, we would stand by you

without fail. But in case we do not succeed, 1 will only request you

to take pity on me. However, I hope any small lapse will not be

noticed by you because you will surely be so engrossed in the main

proceedings.

A20

KEYNOTE ADDRESS

BY

R.K. Garg, Chairman & Managing Director,

Indian Rare Earths Ltd.

1. INTRODUCTION

Uranium is the only primary nuclear fuel and in turn, the only

coamcerlcally worthwhile application of uranium is as nuclear fuel.

Accordingly, with the growth of nuclear power generating capacity the

uraniua industry has grown dramatically over 30 years from virtually no

production in 1950 to around 40,000 T per year by 1980. Apart from its

use *» fuel in power reactors the possibility of its use in nuclear

explosives makes it a material of great stratgic importance and hence it

attracts a number of political and governmental controls. The absence of

bilateral or multilateral safeguard agreements, or other governmental

approvals, therefore prevent certain producer countries from supplying

uranium to some consumer countries.

Though uranium is ubiquitous in nature, rich deposits are rare.

The uranium resources of some of the producing countries of the

world are shown In Table-!. There are many other countries with known

A21reserves of less than 50,000t U but they are not shown In the table.

The Indian resources are Included for comparison. However, I may

add that the production cost for most of the Indian resources would be

well above the $130 per Kg U range, the highest price range upto which

uranium resources in the world are considered.

2. TECHNOLOGY OF URANIUM ORE PROCESSING

The technology of uranium extraction for nuclear applications

usually consists of three steps:

.. the production of marketable concentrate, known as "yellow cake",

analysing about 70% U from the mined ore

.. conversion of this concentrate to a form suitable for final nuclear

fuel and in a purity acceptable for reactor application

.. production of fuel elements to be charged in a reactor

The basic technology for ore processing and production of the

yellow cake was well established by mid 50's to early 60' s. A

simplified flow-sheet which is broadly followed in most of the operating

plants in the world is shown in Figure 1. Though there are many

variations of techniques and types of equipment In use for carrying out

each of the unit operations like sire reduction, leaching, concentration

and final precipitation and recovery of yellow cake, the general flow

sheet has not undergone any profound changes over the years. The only

uranium mill in India working for over two decades follows the same

general technology.

There are two aspects of uranium ore processing technology that need

emphasis. Firstly, in its Initial years of development uranium

hydrometallurgy has freely borrowed the experience of gold and copper

leaching, floculation of leached pulps and solid-liquid separation.

Secondly, the need for processing relatively low grade ores for meeting

the {'rowing uranium demand required special techniques for the

separation and concentration of uranium from the impure and low tenor

leach liquors. This resulted In the Introduction, on a large scale for

A22

Che first tine in the field of matallurgy, of resin ion-exchange in

1950's and of solvent extraction in 1960's. These two techniques have

proved to be extremely versatile and powerful. Both techniques have

later found their way into many hydrometallurgical operations - first in

the nuclear field and subsequently in the non-nuclear field.

3. GROWTH AND PROSPECTS OF URANIUM INDUSTRY

The growth of nuclear power and consequently that of uranium

Industry have not been steady or closely predictable. The earlier

optimistic estimates of nuclear power growth during the 70's have been

revised from time to time.

The world demand for uranium is predicted largely from installed

and projected nuclear power capacity. A number of other factors, of

course, influence this figure. They are the reactor type, efficiency,

degree of fuel enrichment (235U), percentage of 235U in the enrichment

plant tailings, percentage of fuel burn up in the reactor, and whether

the fuel Is reprocessed and the resulting uranium and plutonium are

recycled. The Uranium Institute, London, recently forecast (Table-II)

the uranium needed to fuel existing and planned reactors upto the year

2005. It Is recognized that on a global basis, there are now adequate

resources to meet this demand.

The production of uranium in the past two years and the anticipated

production for 1995 and 2000 is given in Table-Ill. The point to be

noted is that the production In 1987 and 1988 was less than the fuel

needs. This situation Is expected to last till the early part of the

next century. However, there does not seem to be any fear of a real

shortfall developing during this period. The balance fuel requirements

will be met by the users by drawing from the large inventories that were

built up in late 1970's based on optimistic nuclear capacity forecasts.

A second source will be the fuel to be reprocessed from which some

recovered uranium and plutonium are expected to be available for reactor

use.

A23

4. FLUCTUATING URANIUM PRICES

In the short period of 3-4 decades that it existed uranium industry

had a turbulent history. Its growth, as already indicated, though

dramatic, has not been accurately predictable. This is reflected in the

price fluctuation over the years (Fig.2). Following the global oil

crisis in mid 1970's the spot market prices witnessed a steep climb to

US$ 110 per kg U. This was followed by a sharp fall to $50 in 1982.

During 1988 itself, the fall was from $43 to $30 per kg U by the end of

the year. Though the NUEXCO (Nuclear Exchange Corporation) spot prices

do not, for various reasons, reflect the true price paid for

concentrates at any given time, they provide a useful indication of the

prevailing prices. The sharp fall in prices is the direct result of

factoxs like overproduction, lower than predicted demand and the already

large inventories lying with many utility concerns. The prospect of an

immediate or sharp price recovery is viewed in knowledgeable circles as

remote. It apears that the only certainty in the uranium market of the

1980's is its unpredictability. Under these circumstances, it is

reported that some producing countries have also appeared in the market

as buyers, preferlng to buy rather than to produce, to meet their

requirement. However, as earlier mentioned, the sale of uranium is

subject to many governmental or International controls and attractive

price alone cannot be the factor for determining the strategy of

indigeneous production versus procurement from abroad. This is

particularly true of a country like India.

In a situation of declining prices, it is interesting to know how

the major uraniua producers adjusted their strategies. Figure III gives

the production of uraniua during 1980 to 1985. Whereas some countries

have cut down their production, some others have taken measures to

reduce costs. The cost reduction was done not so much by inventing or

adopting revolutionary technologies, except to a small extent, but by

more common sense measures such a»t Increasing cut off grade reducing

capital cost and expenditure for non-essential services, reducing

production wherever possible. In this respect, the Individual measures

A2A

varied from country to country. Some typical cases can be considered

now.

Australia: The cut off grade in the Ranger mine, which is Australia's

biggest and one of the lowest cost uranium mines in the world is 0.5% U.

The present production is 3,000 t U per year which can be boosted to

6,000 t. In the Olympic dam project which has recently started

operation the grade of uranium is only 0.06% U. However, the planned

production is 150,000 t copper, 3,400 kg gold, 23t silver with 3,000 t U

coming as by-product.

U.S.A: In U.S.A. domestic production is drastically cut down on grounds

of economy. A production of 19,000 t U in 1975 has come down by 1988 to

a meagre 2,650 t. To keep down the cost of production a few sand stone

type of deposits are now put on solution aining (also called in situ

leaching (ISL). Significant quantities of uranium (about 1500 t U) are

also produced as by-product from wet process phosphoric acid, from mine

waters and from copper leach solutions.

Technological Improvements have also contributed to the lowering of

uranium production costs. Some of thea are:

. autogenous . or seai-autogen ous grinding of»run-of sine ore (e.g.

sandstone type)

. in situ leaching, wherever the deposit peraits, which eliminates

mining, transporting, grinding and conventional solid-liquid

separation (e.g. sandstone deposits)

. use of high rate thickeners

use of continuous or seal-continuous up-flow ion-exchange

equipaent which eliainate* the need for the costly step of

clarification of leach liquors.

application of ELUEX process which is a combination of

ion-exchange and solvent extraction which peraits the production

of a high grade uranium product.

A25

6. BY-PRODUCTS FROM URANIUM ORES AND URANIUM AS BY-PRODUCT

One way of obtaining low cost uranium is to produce other metals

as by-products from the ore or to obtain uranium as by-product of

other metallurgical operations. There are only a few uranium mines

that have significant payable by-product. However, recovery of

uranium as by-product is a well established practice in some

countries.

Uranium as by-product of gold: In South Africa, the tailings of many

gold ores, after removal .of the precious metal by cyanidation carry

uranium in the range 150-250 ppm. The first full scale plant for

production of uranium concentrates from such tailings was commissioned

in 1952 and by 1957 a total of 17 plants had been erected. Most of

the plants operate on standard sulphuric acid leach, ion-exchange flow

sheet. By 1971, South Africa was producing 3,500-3,800 t U per year

and was one of the important uranium exporters.

Uranium from Phosphoric acid: A very important projected source of by-

product uranium in many parts of the world is the wet process

phosphoric acid (30Z P2 05) which generally carries 60-200 ppm U.

Much attention has been bestowed in U.S.A. on development of a viable

process for recovery of uranium as by-product from this source. The

motivation for this is the fact that the phosphate deposits of that

country contain 4x10 t U. About 30 million tonnes of rock phosphate

is converted into wet-process phosphoric acid annually, setting the

potentially recoverable uranium at 3,000 t per year. After several

years of research in various centres, a very effective solvent

extraction process for producing marketable uranium concentrates from

phosphoric acid has emerged. By 1982 a number of by-product uranium

recovery plants were In operation in U.S.A. with an installed capacity

of about 1,500 t U per year (Table IV). Other countries like Prance,

Belgium, Spain, Yugoslavia and Canada are reported to have set up

plants for the same purpose. Of course, some of these plants are now

reported to be shut down due to lower uranium prices.

A26

Uranium from copper ores; The porphyry copper ores in U.S.A. carry a

small amount of uranium and in all copper leach operations the uranium

finds its way into the final solutions after copper recovery by

cementation. Though the uranium content of these solutions is of the

order of 10 ppm the enormous volumes available make its recovery

attractive. Wyoming Mineral Corp. started in 1977a plant which was

designed to treat about 30,0001 per minute of copper barren solution

by ion-exchange producing 55 t U per year. Anama installed another

plant with a similar capacity in Arizona.

Sea water as a source of uranium; In early 1960's when high rates of

growth of nuclear energy and uranium demand were predicted and it was

feared that long term demands of uranium cannot be met by the then

known reserves attention was directed to the oceans. The ocean water

carrying as much as 4.5x10* t U is the world's largest single source

of uranium though it is present at an extremely low concentration of

3.4 parts per billion. Initial development work on a process to

concentrate uranium from sea water was carried out in U.K. (A.E.R.E)

as a result of which hydrated titanium oxide (HTO) was identified as

an effective absorbent. Subsequently, Federal Republic of Germany and

Japan emerged as Important centres of research in this field. In

addition to HTO, a number of synthetic chelating ion exchange resins

like the polyaaldoxime (PAO) have been Identified as having attractive

absorbing properties. In spite of years of concerted efforts and the

running of a large pilot plant, at considerable cost by a consortium

of industries in Japan, It appears that uranium from the sea can be

obtained only at costs of the order of $600-800 per kg.

7. THE INDIAN SCENE

7.1 Uranium Resources

While the growth of nuclear energy and demand for uranium are

somewhat uncertain in the world, the situation within the country Is

qualitatively different. After weighing the available options for

meeting growth demand of energy within the country, the Government of

India have committed to have an installed nuclear capacity of 10,000

MW(c) based on PHWR by the end of the century. Accordingly, the

Department of Atomic Energy (DAE) worked out a profile for planned

A27

growth of nuclear power (Table V) and various units concerned with the

Implementation of this plan are getting themselves ready for the task.

It Is calculated that for fuelling Initially and for 25 years of

assumed life of the reactors, envisaged to be set up under this plan,

the uranium required is of the order of 40,000 t U, as concentrates.

Against this, the presently known reserves amount to about 60,000 t U

as ore. Taking into account the losses in mining and milling, the

available uranium may meet the requirement. It is also possible that

additional resources will be unveiled in the coming decade. Much of

the ore, however, is of the grade 0.03-0.05% U.0o. It is not feasible

to increase the cut off grade significantly without sacrificing the

available reserves. Hence the production cost of the concentrates may

be of the order of h.3500-5000 per kg U contained. As India is not a

signatory to the NPT, it is not possible to meet the demand by

purchase frost overseas, though the prevailing prices are very

attractive. As a utter of policy, therefore, indigenous production

has to be relied upon.

India is one of the few countries where the entire gamut of

nuclear fuel cycle is well developed and that too entirely by an

indigenous effort. The technology of uranium ore processing is amply

demonstrated by the working of the uranium mill of UCIL at Jaduguda

which has already completed two decades of uninterrupted production.

To meet the growing demand for uranium, UCIL will be opening new mines

and will create additional milling capacity. The annual demand of

uranium by the year 2000 when 10,000 MW (e) installed capacity is

expected to be achieved, is estimated at about 1,500 t.

7.2 By-product Uranium In India

None of the presently known ores have a potential for recovering

economically attractive by-products in a major way which can offset

the high cost of uranium production. However, limited possibilities

exist for by-product uranium. Some of them will be considered now.

Monazlte: Though monazite is relatively rich In uranium (0.30-0.34%)

the total quantity of nineral available from beach s.ind operations is

limited to about 4,500 t per year (likely to increase to about 8,000 t

A28

in the near future). Taking into account the limited demand for

thorium and the problems associated with the chemical processing of

monazite only about 5 to 10 tonnes U per year at present and about 10

to 20 tonnes in future can be expected from this source.

Copper tailings: An attractive source for by-product uranium, though a

poor one, is the tailings from the copper concentration plants in

Singbhum area (Bihar). They carry 80-100 ppm of U Q - At present, a

part of this uranium is recovered as gravity concentrates and

processed in the Jaduguda uranium sill along with the ore from the

mine. In this way, about 302 of the -contained uranium from the

tailings Is recovered. In view of the limited uranium resources of

the country, it is now felt that uraniua recovery can be significantly

improved (to about 70 t per year) If direct chemical leaching is

carried out. This approach is being considered by the department.

Phosphoric acid; As mentioned earlier, wet process phosphoric acid is

considered all over the world as a promising source of by-product

uraniua. A major part of the country's requirement (about 3 million

tonnes) of rock phosphate is met by imports. The phosphoric acid

produced In the country from this raw Material offers the possibility

of recovering uraniua. The know-how for the solvent extraction

process is already available based on the R & D work carried out in

BARC. A proposal to set up the first plant for uraniua recovery

attached to the Cochin plant of FACT is under the active consideration

of DAE.If this is successful, similar plants can be attached to other

phosphoric acid units. At present, soae fertilizer plants (e.g. IFFCO

at Kandla, Madras Fertilizers) depend on imported phosphoric acid

(Merchant grade) of 50-55* P 0. . This is not amenable to solvent

extraction. However, there are soae Indications that in the not too

distant future, these plants aay go in for their own add (30Z PJOJ)

production in which case the total potential for by-product uraniua

from this source can go upto 200 t per year. It Is apparent that

every effort should be put to set up the first plant and prove the

technology as well as economics.

A29

7.3 Possible Reduction of cost of production

Given the grade and capacity of the mines, it appears that a

drastic reduction in cost of uranium production is not possible.

However, some significant reduction can be brought about by:

Increasing the nining capacity as much as possible (say 3,000 t

ore per day or higher)

Improving the overall uranium recovery from the ore beyond the 85%

or so at present obtained

Adopting moving bed ion-exchange (RIP-Resin in pulp) technique

where clarified leach liquors need not be employed.

Recycling major portion of barren liquors after uranium extraction

by IX or SX to leaching circuit, saving on reagent consumption

Combining ore processing and refining steps as much as possible,

avoiding recovery, storage and redissolution of concentrates.

8. CONVERSION PROCESS

So far, the step of obtaining uranium concentrates from the ore

has been considered. The concentrates are too impure to be used in

any nuclear application. 'Conversion' is an Important step in the

nuclear fuel cycle. The main objectives of this operation are:

to convert the uranium ore concentrate into a pure 'Nuclear Grade*

product. Many impurities which are present in the concentrate need to

be reduced to a few parts per million or even fraction of ppm.

to convert the purified product into a suitable chemical form for

the subsequent operation, i.e. fabrication of fuel. The most commonly

used forms are: Uranium metal or UO powder for fuel fabrication and

UF , when the uranium has to be isotopically enriched (235 U) by ao

gaseous diffusion or centrifuge process.

The universally adopted purification process is the one Involving

TBP extraction which takes advantage of the highly selective

extraction of uranium by this solvent. The uranyl nitrate from the

loaded organic is stripped with water and can be converted to

either by denitratlon or by precipitation of ammonium dluranate (ADU)

or ammonium uranyl carbonate (AUC) and calcination. On reduction of

A30

UO , uranium dioxide is obtained which can be converted to oxide fuel

or converted into uranous fluoride. This fluoride, in turn, can be

converted to metallic fuel by metallothermic reduction (using

magnesium) or to UF for isotopic separation. The lsotopically

enriched UF, can be hydrolysed with water, precipitated as ADU and

fB—itpi*m*m4 «• M V mm# converted into UO- for fuel fabrication

(Fig.3).

There are five major refining plants in the western world (Table

VI). Although the process used in these plants is about the same, the

equipment is different, e.g. while the BNFL uses mixer-settlers,

Coaurhex use agitated columns and Eldorado Nuclear a combination of

Mixco columns and pulse columns. It is generally believed that

conversion plants require an annual production level of 5,000 t U to

be economic.

For production of U02 and UF^, rotary furnaces and fluidised bed

reactors are In common use. For the production of UF gfluidised bed

reactors and flame reactors are being used.

In this country, a refining unit with a capacity to produce 25 t

uranium metal per year was set up mm fat back as 1959. Its capacity

has been Increased recently to meet additional requirement of fuel for

the DHRUVA reactor. The refineries set up' so far are of small

capacity 100-200 t U per year but with future demand in view higher

capacities upto 500 t are being planned. For solvent extraction, we

have experience of both pulse columns and mixer-settlers. In

addition, a very significant contribution in this area has been the

development of the slurry extractor at the Nuclear Fuel Complex. The

slurry obtained after digestion of the yellow cake with nitric acid

can be directly fed to the extractor without putting it through the

difficult step of filtration and washing. The experience with this

extractor for the past 2-3 years has been very encouraging.

9. CONCLUSION

In conclusion, it can be said that in spite of some set back in

A31

the rate of nuclear power growth, in the world, future will see only a

net increase in the installed capacity. Consequently, the demand for

uranium is expected to grow steadily. The presently known resources

in the world are sufficient to meet the demand for the foreseeable

future. At present the installed ore processing and refining capacity

is more than the current demand. Hence only a slow growth of

additional capacity In these areas is foreseen. Due to slack in demand

and heavy inventories, the price of uranium concentrates has steadily

fallen in the recent past. The trend may not be reversed in the next

few years.

In India, a committed programme for increasing installed nuclear

power capacity to 10,000 MW (e) by the end of the century has

necessitated a rapid growth of uranium mining and milling capacities.

The known reserves are just sufficient for the planned growth but

there is a need for stepping up exploration and identifying additional

resources, possibly of higher grade. Meanwhile, the factors which

need attention are:

. bringing the deposits into production as early as possible

reducing the cost of production

improving recoveries in Billing and conversion plants

. Improving recoveries in fuel production

. since the scale of operations in all parts of the fuel cycle will

be increased several fold compared to the present level, measures

for tackling environmental problems associated with each step should

be worked out carefully. Greater mechanisation will also be necessary

to reduce manual handling and consequently radiation exposure.

improving recovery from copper tailings and take steps for

incorporating uranium recovery circuits in phosphoric acid plants.

I hope the details pertaining to some of the aspects covered in

my talk will be forthcoming from the series of invited talks and

technical presentations that will be heard during this symposium.

URANIUM

Country

Australia

Brazil

Canada

France

India

Namibia

Nigeria

South Africa

U.S.A.

Others

A32

TABLE - I

RESOURCES OF MAJOR PRODUCING COUNTRIES

Data * as on 1.1.1981,

Reasonably Assured

•000 t U

317

119

258

74.9

32

135

160

356

605

237

cost range US $ 130/kg U

Estimated Additional

•000 t U

285

81

760

46.5

25

53

53

175

1,095

147

Total 2,2293 2,720

* Only for countries outside the centrally planned economies

Source: Joint report by the OECD Nuclear Energy Agency and the IAEA,

1983.

A33

TABLE II

The Uranium needed to fuel Reactors ('000 t U)

1988 1989 1990 1995 2000 2005

1986 forecast 44.0 44.4 44.5 49.4 52.3 na

1988 forecast 42.7 44.1 47.2 51.0 55.0 56.0

Ref: Metals & Minerals Annual Review - 1989

A34

TABLE III

Uranlua Production In the World ('000 t U)

Country 1987

Australia

Canada

Europe (MainlyFrance)

Naalbla

Gabon Niger

(Central Africa)

South Africa

U.S.A

Others

3.8

12.4

3.7

3.5

3.8

4.0

4.8

0.7

1988

3.6

12.4

3.8

3.5

3.9

3.8

5.2

0.6

Estimated

1995

8.7

14

3.

4.

4.

0.

5.

5.

.8

1

0

5

77

7

6

2000

9.9

17

1.

3.

2.

0.

4.

6.

.7

4

5

8

77

9

4

Total 36.7 36.8 47.3 47.4

Ref: Mining Annual Review, 1989

Metals and Minerals Annual Review 1989

A35

TABLE IV

Plants in U.S.A. For By-Product Uranium Recovery From Phosphoric acid

Capacity t/y

Company Location P 02 5

Free Port Uraniua Recovery Louisiane 6,80,000 265Co.

Wyoaing Mineral Corp.

Gardialr Incorp.

International Minerals&Cheaicals Corp.

Earth Sciences, Inc.

FloridaM

H

•t

Alberta, Canada

3,60,000

4,50,000

5,00,000

7,60,000

1,190,000

145,000

135

160

170

290

485

40

Source : Uraniua Institute, London

International Conference, 1983.

A36

TABLE V

Planned Growth of Installed Nuclear Energy Capacity In India

By the end of Total capacity (Cumulative)

MW (e)

7th Five Year Plan 1465

6th Five Year Plan 2170

9th Five Year Plan 8550

2000-2001 10,050

A37

TABLE VI

Major Uranium Refineries in the World

CapacityPlant „,

t.U/year

Allied Chemicals (U.S.A.) 12,700

BNFL (U.K.) 9,500

Coaurhex (France) 12,000

Eldorado (Canada) 9,000

Sequoyat Fuels (U.S.A.) 9,090

Total Western World 52,300

A38

URANIUM MINING JN INDIA

PAST. PRESENT AND FUTURE

M.K. BATRA

INTRODUCTION

The search to locate indigenous sources of uranium began

as a sequel to the decision to harness atomic energy for indus-

trial purposes in the country. A raw materials division was set

up and temas of Geologists started exploration in the various

parts of the Country. The areas which were considered likely to

have uranium occux^tnces included Singhbhum Thurst Zone in South

Bihar. The area was already known for its copper sulphide miner-

alization, and operating copper mines were located therein.

Association of copper and uranium had been reported in many parts

of the world, though no commercial deposit had yet been found. A

sample of uranium had been picked up by a prospector from one of

the copper mines as early as in 1937. The sample had been analy-

sed to contain uranium, in the laboratories of G.S.I, at Calcut-

ta. In 1950, therefore close examination of this 160 Km. long

mineral zone, out cropping on the ridge of a hill, which could

have a sizeable potential was revealed at Jaduguda. This turned

out to be a major deposit and has remained the best located so

far.

In this belt, series of rock formations have been strongly

folded and highly metamorphised. A constant techtonic movement

hus created a zone of t.hurst. The rocks towards the North have

been bodily thrown against the rocks towards South. The zone of

A39

thursting had been completely sheared and became a favourable

place for deposition of mineralized solutions. It is1 in this

zone of sheared rocks that Uranium, Copper, Nickle and Molybde-

num joineralisation has taken place. There had been two phases of

mineralisation; a high temperature oxide phase and later, a low

temperature sulphide phase. In the oxide phase, minerals such as

apatite, magnetite and uranium were deposited, while in the

sulphide phase, minerals of copper, nickle and molybdenum were

formed. The age of mineralisation is stated to be about 1000

million years. Importance of associate rocks and minerals lies

in fact that these often lead to 'finding of principle mineral.

RETHINKING:

Recently, there has been re-thinking in mode of deposi-

tion, in this area, though views to the contrary have been aired

from time to time. A school of Geologists are of the view that

the area is of sedimentary origin.and the quartz pebble conglom-

erate formed in the thurst zone are from a river bed. If this

theory holds true, there is a great possibility of existance of

wide and better ore zone towards North of the present working

harisons of both uranium and copper. A programme to test this

possibility has been drawn by AMD and a test bore hole is now in

progress.

ORE BODY AI JADUGUDA

The choice of mining method is normally dictated by the

A40

characteristics of the ore body. The ore body at Jaduguda is

lenticular in shape. The lenses pinch and swell and the width

varies from a few centimeters to 5-6 meters. The lenses are

separated by waste patches. The dip of the ore body on average

is about 45 degrees, but the veins take a roll, as they go in

depth and become very flat. This erratic behaviour has hampered

adoption of more efficient and high productive mining method. A

mining method caled 'Cut and Fill' had to be edopted to ensure

controlled breakage, with a view to eliminate high dilution from

waste rock. This method, of course, helps ensure safety in

mining operations as wall rocks in the thurst zone are highly

jointed and tend to break loose without *uch warning. The fill-

ing system ensures ground stability.

ECONOMICS;

While the uraniua Mineralistion is wide spread in the

ar*a, the economic length of the ore body at Jaduguda is only

about 850 meters. During exploration in the fifties, the ore

body was traced out to a depth of about 450 meters by diamond

drilling and about 4.5 million tons of ore reserves at an average

grade of 0.065% e U3O8 were established. Later bore holes proved

the continuity to about 850 meters. A few still deeper bore

holes have inter-sected the ore body and found it to be still

persisting. A copper mine, in the neighbour-hood, at Mosaboni,

has workings at a depth of about 5000 feet, at present and there

A41

is no reason, why the uranium lodes should not go this far and

still further.

Along with diamond drilling, exploratory mining was also

carried out at Jaduguda. Adits were driven on the face and in

the foot of the hills and levels were driven along the ore body.

This gave sufficient information about the rock type, ore horizon

and ore characteristics. Opening of the ore body provided bulk

samples for carrying out metallurgical tests.

As compared to presence of mineralization; a deposit is

called an ore deposit when the mineral is present in sufficient

quantities and in quality to justify an adequate pay back period

and adequate return of investment. A mine is a wasting asset. A

large capital is to be invested, to begin with, to set up facit-

lities for mining and processing of ore and then the depletion

starts. Great caution is therefore necessary to make estimate of

ore reserves so that the investment does not come to grief.

There was quite a hesitation in taking up Jaduguda deposit

for commercial exploitation. The ore body was small, the grade

of ore was not high enough to be excited about, and underground

method of mining, the only alternative in this case, was not

conducive enough for high production rate. About this time,

number of large deposits were being discovered in the Western

world, in fact extensive uranium fields, like Elliot Lake in

Canada, Ambrosiu area in New Mexico and very high grade intrusive

A42

deposits in Colorado in U.S.A., and in Gaban, Niger & Nambibia.

Economic studies showed Jaduguda in poor light when comparisons

were made. Department of Atomic Energy had invited teams from

internationally known mining companies, like Rio Tinto Zinc and

later from Prance to evaluate the deposit and their reports were

not too encouraging and implied that there was not enough econom-

ic justification for opening of Jaduguda when uranium could be

had in abundance from then.

THE DECISION

However by 1961, decision was taken to open up the depos-

it and to set up a mine and a mill. Work had proceeded ahead at

Trombay in drawing of the process flow sheet. That year, Jadugu-

da Mines Project was set up to concentrate efforts on developing

the nine. Decision was taken to sink a shaft in middle of the

ore body to provide an acceessway for hoisting of the ore and

for the horizons to be developed. This work assumed priority as

the mill was being constructed,, simultaneously which would be

ready earlier. Stoping work was therefore, taken up in the

levels, which had been developed through the adits and which

would provide stock piles of ore for the mill till the commence-

ment of regular productin from the mine. It was also decided to

sink the shaft in two phases, so that the mine could be brought

into production earlier. The first phase consisted from surface

to a depth of 315 meters.

A43

UCIL:

In 1967, the two projects, Jaduguda Mines project and

Uranium Mill Project were merged and a Public Sector Company,

UCIL. under the administrative control of Department of Atomic

Energy was formed, with a specific objective of mining and mill-

ing of uranium ore in the country. By 1968, shaft along with the

ore pass system, underground loading and crushing stations were

made ready to produce 1000 tons of ore per day. The mill went

into production a little earlier.

Ilnd stage shaft sinking when the shaft was deepened from

315 meters to 640 meters was carried out, along with the produc-

tion of ore from the top levels. A noval method was used for

shaft construction. The main ore pass was sunk and the shaft was

raised from bottom to top. Instances of use of this method are

very few and far between in the world. The Ilnd stage was com-

pleted in 1977 and mining was commenced in the deeper levels.

The shaft is now being taken up in Illrd stage now, where an

auxiliary shaft is being sunk from 555 meters level to 850 meter

level. This will allow the mine to continue production till the

end of this century. As the ore body is still open, mining is

likely to continue further down.

In the earlier stages, to boost up production and to build

up ore stocks, shrinkage system, where bulk of the ore could be

left in the stopes for drawl later on and open timbered methods

A44

of stoping were used. Shrinkage stopes provided opportunity of

application of solution mining. Due to flat dip, some ore was

adhering to foot wall; even after drawl from the chutes. Barren

solution from the mill was sprinkled into the stopes and re-

circulted till values were built up This water rich in uranium,

was then pumped to the aill for uranium extraction. This contin-

ued for quite so«e time and considerable expertise has been built

up in this regard. While stringers are difficult to 'leach,

finely broken ore can be leached reasonably. Later, cut and fill

method was standardised; the voids created by mining are filled

up with dislimed mill tailings. You will hear about these sys-

tems and see some slides in the papers being presented later on.

Jaduguda was almost the first underground metal mine to go

into production after independence. In keeping with the stand-

ards of the Atomic Energy Establishments many new technologies

were used and for the first time in India, a concrete tower, to

house friction type winders at the top was built with slip form

technique brought from Sweden. The system was so well liked and

absorbed that it was later used for lining of the shaft with

concrete. The equipment brought from Sweden on rental basis, was

purchased and retained. Alimak raise climbing equipment was used

for driving of raises with speed and safety. Tyre mounted load-

ers were introduced underground for the first time in India, for

handling of broken ore. Since then, use of slip form arid load,

A45

haul and dump loaders have been widely used in the Country.

Grouted Rock Bolting is now used extensively as a support system

underground and this has replaced timber supports.

Jaduguda is well designed mine and has good functional lay

outs, not only for production purposes but also for transporta-

tion (use of diesel locomotives) and drainage system. Stope

wagons have been used for upper drilling and prilled ammonium

nitrate is used for blasting of ore. Out of about ore reserves

of 10.5 Million tons, upto 555 Meters depth, about 4.5m tonnes

have been extracted so far.

Hilling is an integral part of a metal mine. In ore

processing, also, new technologies were used in extensive manner

as part of hydro-Metallurgy, like leaching of ores in Pachukas,

use of drum filters, ion-exchange system and re-precipitation

techniques.

THE PRESENT SETTING; BHATIN;

A new Mine has since been opened at Bhatin, about 3 KM.

froM Jaduguda. The ore froM this Mine is brought by duMpers to

Jaduguda Mill for processing. Ore reserves here upto a depth of

about 500 Meters total to about 2.5 Million tons of ore, at a

grade of .045%. The production rating of this mine ia about 250

tons per day. Opened in 1987, designs of this Mine were prepared

in the Corporation itself.

A46

URANIUM RECOVERY PLANTS:

An auxiliary source of uranium has been copper tailings.

The copper ores of Singhbhum contain small values of Uranium and

these are separated from the copper tailings by gravity separa-

tion method. The recovery plants are located adjacent to the

three copper concentrators. The feed grade varies from .008%

to .01% and upgradation is about 10 times. The mineral concen-

trate with grades of 0.08X to 0.1% are transported to Jaduguda

and mixed with ore for extraction of uranium. This has been a

good source of uranium. To improve recovery, chemical treatment

employing low acid leach is being considered. As copper tailings

after recovery of uranium may have still manganese pollution, use

of bacteria in place of manganese as an oxident is being studied.

Simultaneously, use of sliae tables 'to recover uranium now

going out as ultra fine particles is being studied. Success of

these studies will establish this source on.more firm basis.

EXPANSION i MILL

The Jaduguda mill was expanded two years ago and is now

capable of handling about 1400 tons of ore per day, increased

feed coming from Bhatin nine and the uraniun recovery plants.

JBI£ PRODUCTS:

Another distinct feature of Jaduguda is recovery of acces

A4 7

sory minerals occuring with the ore. In the Bye products Recov-

ery Planti copper molybdenum and magnetite are recovered while

copper is. recovered before extraction of uranium, magnetite is

extracted from the tailings. This has been a notable achievement

as otherwise these values would have been lost in the tailings.

The operations are profitable and make a handpome contribution.

Sometimes economics of mining of principle minerals itself is

decided by presence of bye products.

FUTURE Q£ MINING;

For 10,000 MW programme, requirement of the concentrate

is estimated at about 1350 tons per annum. Constant review is

therefore, required to be made for opening of new deposits, which

have been explored by AMD.

Two such deposits taken up presently for construction are

Narwapahar and Turamdih East, where a mine each will be set up

with a production capacity of 1500 tons of ore per day, and a

mill at Turamdih to treat ore from both the mines i.e. 3000 tons

per day. Ore from Nnrwapahar mine will be transported by an

aerial rope-wuy. These deposits are located at a distance of 12

Km. and 25 Km. from Jaduguda respectively.

Both will be underground mines. In keeping with the

latest trends, these will be trackless mines, access to the ore

body will be through declines rumps, stopping at about 10 de-

grees. Ore will be huiiled by low profile dumpers. Men will

A48

travel to the working places in passenger carriers. Bulk mining

methods like post pillar for wide ore bodies, more than 6 meters

wide, room and pillar for narrow lenses have been proposed for

the mine. Higher productivity levels have been earmarked for

these mines. This has been an economic necessity, as grade of

the mines is lower to that at Jaduguda. Both the deposits have

reserves of about 10 Billion tons each, with grade of 0.058X at

Narwapahar and 0.045% at Turamdih.

The major improvement will be in ventilation system. The

entry system and working methods are such as to'provide fresh air

directly to each face; unlike in shaft system, where some air

does get re-circulated. There has been a considerable lowering

down of international standards with regard to radon concentra-

tion and the up-dated ventilation system will help achieve the

rigid standards.

In the new mill too, losses are likely to be reduced with

introduction of horizontal belt filters and continuous counter

current, fluidised bed Ion-exchange system. Tailing disposal

system has also an improved design about which you will l«arn

from a paper being presented later.

Due to high capital costs involved in the projects and

nature of underground mining of low grade ore production cost of

uranium concentrates is estimated to be quite high Cost effec-

A49

tiveness, will therefore, be a paramount requirement. Our expe-

rience at Jaduguda has been, that while costs of mining and

milling per ton of ore have been quite competitive, inspite of

high costs of some inputs, cost per Kg. of concentrate obtained

comes higher, because of lower tenor of ore.

We have yet not been able to discover a large high grade

deposit or deposits and therefore must resort to small deposits

of comparatively low grade, most of which occur in Singhbhum

district. These deposits offer good possibility of adoption of

solution mining and heap leaching techniques. We have carried

out good amount of work in these fields on experimental scale and

time has come when such techniques must receive good impetus.

Preg. liquor obtained at sites can be transported in tankers to

central mills at Jaduguda or Turamdih. It may be pointed out

here that about 700 tons of uranium is produced in the world by

such methods out of total production of about 42,000 tons. At

Denison Mine in Canada, about 18% of the mine production comes

from mine water. Adoption of such approach for us will shorten

the pay back period and the start of production can be short

enough to wait for the discovery of richer deposits. To meet the

target production, resort to such technique is a must. This work

can be undertaken in shorter time and on a lower investment and

can be tailored off when rich grade deposits are found.

A50

OPEN CAST;

Underground mining is very restrictive in nature. A

deposit where open pit mining can be practised carries number of

advantages, in, fast start up, higher production, cost reduction,

computerised control etc., The deposits located at Doraia Sat in

Meghalaya and Turamdih West in Singhbhua have very favourable

stripping ratios. The advantages at Doaia Sat is much more as

the ore here is of sand stone type and it should be possible to

heap leach effectively, lower grade ore removed from the top

layers; remaining being treated in a conventional mill.

Because of the wider range avilable here, ore sorting

machines, based on gamma ray emissions, can be used to separate

lower grade ore. In open cast area, number of land reclamation

measures have now been devised. In some case, it has even been

possible to upgrade the land. Solutions are available therefore

in this regard.

However, Domia Sat has a handicap of difficult logistics.\

Considering vastness of the source, these must be overcome.

TAILING DISPOSAL:

Anotther area which poses a stiff challenge and must be

tackled effectively is desposal of tailings. Even future of some

deposits will be decided on this account. Pressure on land

requirement must be reduced alongwith the measures taken for

environmental control. In cut and fill method, only coarse

A51

fraction of the tailings can be used, which is hardly 40X of the

total. Therefore 60% must be impounded on surface, requiring a

large land area. Use of tailings can be increased by adopting

mining methods where delayed filling can be used. Replacement of

hydraulic filling with pnematic stowing can be a possibility.

These tailings can also be agglomerated. Considerable work needs

to be done to bring these concepts to practical applications.

SHORTENING THE GAP:

Considerable time elapses now, between preparation of DPR

and commencement of actual work on ground. Future projects can

ill afford this delay. A close collaboration is necessary,

therefore, between both, exploration and exploitation agencies.

Some work of conceptual in nature can be taken up early if a

deposit in exploration is showing a promise. Some pre-

feasibility studies on provision and scale of infrastructure

facilities can be undertaken simultaneously. Statutory clearance

also take time and need speeding up.

LAND ACQUISITION. REHABILITATION AND RECLAMATION:

Acquisition of some land for opening up a mine and a plant

may be inevitable and must be kept to the minimum. Acquisition

is a time consuming process and action is to be initiated early

enough to uvoid delays. Rehabilitation of displaced persons is a

social responsibility und hus to be taken up earnestly. Skills

A52

may have to be imparted to displaced persons for them to be

gainfully employed in the projects/ A good expertise is avail-

able, at present, for reclaimation of land, ravaged by mining,

particularly by open cast operations. Not only it is possible to

reclaim the land, but it is also possible to upgrade the same.

Mention must be made here regarding requirement of environment

management particularly of liquid effluents, both from the mine

and the plant.

COST REDUCTION:

Future of mining lies in competitiveness and system there-

fore, must incorporate cost reduction provisions. Mining meth-

ods, where the operations can be carried out independently and

not in cyclic order as in the cut and fill method will be more

useful.

While high degree of mechanisation,does not necessarily

mean cost reduction, there are certain aspects in mining opera-

tions where closer look is required. One such area is drilling.

Our costs in drilling and blasting are very high Mechanisation

of drilling operations and use of hydraulic drlling equipment may

be the answer. Time has come that use of Raise Borer needs to be

considered in depth. This type of equipment can eliminate the

delays involved in developing a mine.

During opening of Juduguda, we took lead in many ureas

of underground mining. This has paid us handsome dividends. We

A53

have to be prepared once again to blaze a trail. Mining and

milling of low grade ores has its problems which must be faced.

In short, mining techniques, in future, will have to

undergo a drastic change. There are pressures enough for that.

While we here are presently engaged in construction of

Narwapahar and Turamdih, Cigar Lake mine is being prepared for

production in Canada. Fro* 3000 tons of ore per day, we will be

getting about 320 tons of U3O8 per annum. Cigar Lake will be a

100 tons per day proposition and production of concentrates is

estimated at 4,200 tons. Very attractive and exciting indeed,

but then Mining of high grade deposits can have problems of its

own.

I am thankful to the organizers for giving me an opportuni-

ty to present before you, a birds' eye view of'the scenario here

at home.

Thank you,

SESSION I I A

U R A H I V M P R O S P E C T I N G

Chairman : SHRI S.3A3TRYChief Geologist UCIL.

STRUCTURE AS A GUIDE FOR URANIUM EXPLORATION IN THE TURAMDIHMOHULDIH AREA, SINGHBHUM DISTRICT, BIHAR

R . MOHANTY and M . B . VERMAAtomic Minerals Division

34, Khasmahal, Tatanagar - 631 002

Uranium mineralisation at Turamdih i s hosted by chlorite-quartz schist±apatite and magnetite* whereas at Mohuldih, i toccurs in the immediately underlying quartzite and tourmaline-bearing sericite schist. The ore horizons are in the form ofa number of lodes, concordant with the schistosity of the hostrocks, and separated from each other by a few tens of metresof poorly-mineralised or barren rocks.

Of the three deformation episodes (F^ F^ and F,) deci-pherable in the area, evaluation drilling and structural ana-lysis reveal that the subsurface behaviour of the ore body i smostly affected by the F2 fold movement. Critical informationon such structural guides for mineralisation will help in pla-nning evaluation drilling programmes in the contiguous area tosubstantially augment the presently-known reserves of uranium*

- 2 -

INTRODUCTION

The Singhbhum Shear Zone (SSZ) i s w e l l known f o r i t s Cu-U

mineralisation. Though i t extends for about 200 km# only the

eastern 100 km contains the major uranium and copper deposits.

The western portion of this eastern stretch constitutes the

Turamdih - Mohuldih sector which probably houses the largest

uranium deposit of the belt. This sector with an area of 5 km x

2 km, l i es within 10 km from Tatanagar (Fig. l ) . in this paper

we have attempted to discuss the exploration stages for uran-

ium and the effect of structure on the sub-surface behaviour

of the ore body in the Turamdih - Mohuldih area*

GEOLOGY AND LOCAL STRUCTUR1

Pioneering works on geology and structure of the SSZ,

among others, include those of Sunn (1940)* Dunn and Dsy (1942);

Sarkar and Sah«, (1962); Sarkar (1964); and Mukhopadhyay (1976,

1984)• The Turamdih - Mohuldih sector exposes sodagranite

underlying a mstasedinentary sequence comprising banded magnetite

ouartzlte, serldte schist and chlorite schist belonging to the

Iron Ore Stag* (Dunn and Day, 1942) or the Dhalbhum Stag* (Sarkar

and Sana* 1962). This rone is bounded by the Dhanjorl Formation

in the south and the mica schists of the Chalbasa stage in the

north (Fig l), and has bean referred to as the Shear Zone* in

the centra of which the Mohuldih - Turamdih area lies. The

effect of shearing is most intense In the central part, which

gradually decreases in intensity both towards north and 'south*

The uranium and copper mineralisations are mostly asso-

ciated with the rocks in the aforesaid shear zone. Momm of the

promising uranium occurrences along the SSZ, with which the

Atomic Minerals Division Is presently Involved and their host

rocks are summarised in Table-I*

- 3 -Table .

Rock type

N Game t i ferous^ mica schist and

guartzltes

Chlorite schist+ apatite and

magnetite

. Sericite schist ±tourmalineBanded magnetitequartzite

S Soda granite

I

Stratigraphy asreferred byDunn and Dey,1942

Chaibasa Stage

Iron Ore Stage

.do-

Soda granite

Known uraniumoccurrences

Bagjata*,Kanyaluka,Gohala

Narwapahar*Turamdih*Garadih*Rajgaon

Mohuldih*Bangurdlh

* economically viable deposits

, The local structure i s in no way different from the regionalstructure described by Hikhopedhyay (1964). As described by him*there are three folding episodes decipherable in the area* In abroad sense, the earliest deformation (Fj) i s of tight to iso-clinal reclined folds with the development of a pervasive axialplane shlstoslty (*x) which, at most places, i s parallel to thebedding (So). The general trend of the foliations i s MM# - SSIdipping 30-40° towards HI. The hinge zones, where the 8Q and S^are supposed to be perpendicular to each other, are, however,hard to find* The Fj folds are so much drawn out and affactedby later extensive mylonltisatlon that small scale Fj folds havebecome scarce* The down dip llneatlons which are profuselydeveloped on S planes parallel the Fj fold axis and hence areTx lineatlons (L1). Incidentally, these lineatlons also parallelthe strlatlon lineations (a-llneatlons) pertaining to the laterphase of folding* However, F folds in the mappable seal* aresometimes preserved In the quartzite outcrops (rig.2)*

- 4 -

The second generation of folds (F ) trends ESB-WNW and arenonplunging to low plunging either towards east or west. Mostof the small scale folds v is ible on the surface belong to thisgeneration. The earlier schistosity S. has been affected bythis folding. A se t of crenulation cleavage Cs

2^ Parallel toi t s axial plane has been developed more dominantly in the sch i s -tose rocks. The lineations (L_) pertaining to F~ folds occurin the form of puckers on the S surface. Not much variationin attitude of F~ folds i s seen because the S- surfaces arefairly consistent in their attitude and F1 hinges are very rare.

Overprinted on them are the P^ folds i n t n e form of broadwarps with very high wavelength/amplitude rat io . The axes ofthese folds trend almost N-S with moderate amount of plungetowards north. No small scale manifestTTations are, however,recognisable except minor strike swings. The e f fec t of thesethree generations of folds on the ore body in the subsurface arediscussed in the following.

URANIUM MINERALISATION

Host rocks

The uranium occurrence, at Mohuldih and Turatndih was knownduring la te f i f t i e s (Bhola, 1965). Mineralisation at Turamdihi s hosted by chlorite quartz schist ± apatite and magnetitewhereas that at Mohuldih i s hosted by the immediately underlyingunit of s e r i c i t e schist and banded magnetite quartz!te (Table-I).The chlorite schis t at Mohuldih, which i s in the strike continua-tion of 'Airamdih, however, contains impersistent uranium horizons.

Exploration

Exploration by evaluation dril l ing at Turamdih and Mohuldihhas been carried out through various stages during the l a s t threedecades. At Turamdih area* the mineralisation occurs over 1.5 km

strike length with approximately 1 km plan width (Fig .3 ) . Since

- 5 -

the structure has played a great role in transposing the radio-

active bands both in the surface and the subsurface zones, explo-

ration had been undertaken in different blocks, such as, Turamdih

East, Nandup, Turamdih North, and Turamdih South (Fig.3) at diffe-

rent time6. However, after detailed exploration i t i s now under-

stood that the ore bodies of these blocks are manifestations of

one and the same body, affected by all the three deformations

resulting in i t s occurrence at different levels and in different

shapes. The evaluation drilling at these blocks was done at a

grid interval of SO to 60 m along 6trike and 100 to 120 m along

dip. At Turamdih North, however, the dip interval has been

brought down to 50 to 60 m. These Intervals, both along strike

and dip, have been decided not by any statistical calculations,

but by trial and error to Maintain the variation in behaviour of

the ore body to the minimum. It can be mentioned here that the

outcrop of the ore body at Nandup continues below the surface at

Turamdih South and Turamdih Cast only to crop out again to the

north at Turamdih North and Keruadungrl.

At Mohuldih which l i e s about 2 km west of Turamdih, minera-

lisation occurs on surface over 350 m strike length with sub-

surface continuity of l i t t l e more than 1 km. Exploration at

Mohuldih has been done In 2 stages - once in 1969-70 and next

during 1982-87. Drilling was done at an interval of 60 m along

strike and 120 m along the dip.

Correlation studies and sub-surface structure

Uranium lodes along the Shear Zone are basically controlled

by stratigraphy0 llthology and geochemistry (Rao and Rao, 1983)

on which structural effects are superimposed. I t i s of interest

to know how each of the three folding episodes described earlier,

has affected the uranium lodes in the subsurface. While the

mineralisation i s confined to one particular l i thic unit, i t

occurs in the form of a number of layers and i s folded sympathe-

tically with the host*

Since the Fj folds are isoclinally reclined in nature and

-6-

plunging towards NE, their impressions are better seen along

the strike sections (trending ESE-WNW) of the ore body (Fig, 4a,

4b). In these sections, the ore body closes either towards east

or west. Such closures of very small dimension of few tens of

metre are observed. Because of the F^ fold, the lodes are repea-

ted to form a number of horizons. At the hinge areas of such

closures, thicker mineralisation is normally intercepted as in

case of Turamdih (Fig. 4a). In Mohuldih, mineralisation occurs

in the form of two prominent lodes, the gap between which narrows

down, to finally coalesce in the western end and widens to about

40 metres at the eastern extremity (Fig. 5a, 5b)• Such coalescing

zones or perfect closures pertain to the F^ folding movement.

Sometimes high angle relationship between the bedding and the

schistosity is observed along the cores at such closures.

Of the three folding episodes, the effect of F2 folding on

the ore body and the formations is maximum. These folds trend

WNW-ESE with their axial planes steeply dipping towards NE. Their

northern limbs are always longer than the southern limbs. Many

times, these folds are intercepted along the boreholes (Fig. 6),

and therefore, care has been taken to consider the true thickness

and not the apparent thickness thus intercepted. The correlation

of the ore bands has, accordingly, been dons taking these folds

into consideration* The manifestation of these folds is best seen

along the dip sections (Pig. 7 and 8 ) . At Turamdih* it is this

F2 fold which helps in linking the deposits of Nandup, Turamdih

South and Turamdih North with each other and establishing the

ore body as one and the same. In Fig.8, the southern portion,

where the mineralisation is at or very near to the surface* is

the Nandup deposit. This ore zone goes below the surface for a

plan width of 500 metres to form the Turamdih South deposit.

This, with the help of a large F^ synclinal fold, surfaces again

to the north resulting in the Turamdih North deposit. In Turam-

dih South, the frequency of F2 fold increases.so that the dip

of the enveloping surface (imaginary surface joining crest to

crest of the folds) becomes horizontal to sub-horizontal resul-

ting in shallow interception of the lodes even in the downdlp

- 7 -

dlrection. Since the F2 fold axes are horizontal to subhori-

zontal* the lodes are intercepted at almost same level in any

strike direction. At Mohuldih, however, (Fig.7) not only the

dip of the enveloping surface becomes subhorizontal at certain

depth but also a gradient is observed even along the strike

direction (Fig. 5b) beyond the 5th series of exploration. This

could be due to acute angle relationship between F and F. axes

The interference of F , F_ and the later F. folds brings out an

interesting subsurface mosaic. The true dips, calculated from

the apparent dips based on the marker intercept e.g. contact of

schists and quartzite indicates a gradual change in strike from

NW-SE along 1st series to NE-SW along 9th series (Table-II) of

boreholes*

Table II

S.NO.

1 .

2 .

3 .4*

5.6 .

7 .

8 .

Series

I

I I

I I I

IV

V

VI

VII

VIII

and II

and III

and IV

and V

and VI

and VII

and VIII

and IX

Dip amount and direction

30° towards N30°E

30° towards N35°E

20° towards M40°E15° towards N70 B28° towards N80°I24° towards S70°S26° toward* S70°E30° towards S50°«

The structural contour drawn for the contact as well as the

mid point of the ore body (Fig.9) also corroborates these changes

along strike in the subsurface horizons.

The third deformation el episode <F3) has, however, the least

affect on the or* body* Since the F3 axas are subparallcl to F

axes, minor warps are observed in the ore body along the strike

sections (Fig* 4)«

- 8 -

The above discussion, thus, reveals that the uranium minera-

lisation is basically lithic-controlled and predeformational. The

F, folds have caused the repetition of ore horizons, whereas the

F2 deformation has contributed in bringing the lodes to shallow

levels, at places even to the surface.

Substantial reserves of uranium have been proved in the

Turaindih-Mohuldih sector. Additional reserves will be proved in

future, in areas like Keruadungri (adjacent to the Turamdih North)

and the intervening gap between Turamdih and Mohuldih, where explo-

ration by drilling i s being carried out at present. Detailed struc-

tural studies, as done in the case of Turamdih-Mohuldih, will go a

long way in planning exploration strategies in these areas also.

ACKN OWLEDGEMENT

The authors are greatly indebted to Shri A.C. Saraswat,

Director, AMD for his encouragement in writing this paper. The

constant guidance by Shri K.K.Slnha, Regional Director* Eastern

Region and Shri S.C.Verma, Project Manager, and involvement at

every stage by Shri L.D. Upadhyay, Deputy Project Manager are

thankfully acknowledged. Thanks are also due to the previous

workers of the AMD especially S/Shri K.D. Agarwal, R.M.Sinha,

R.K. Gupta and E.U. Khan whose unpublished reports have formed a

base for this study, and to Shri H.M.Verma and R. Dhana Raju for

crit ical ly reviewing the paper*

REFERENCES

Bhola, K.L. (1965) » Radioact ive d e p o s i t s i n I n d i a . In 'Uranium

Prospecting and mining i n I n d i a 1 , D.A.E. , Jaduguda,

p.1-41.

Dunn, J.A. (1940) : The stratigraphy of South Singhbhum, Mem. Geol.

Surv. India, V.63 (3),

Dunn, J.A. and Dey, A.K. (1942) * Geology and petrology of Eastern

Singhbhum and surrounding areas. Mem* Geol. Survey*

India, V*69(2).

- 9 -

Mukhopadhyay, D. (1976) t Precambrian stratigraphy of Singhbhum -

the problems and a prospect. Ind. Jour, Earth. Sc i . ,

V.3, p.208-219.

Mukhopadhyay, D, (1984) * The Singhbhum Shear Zone and i t s place

in the evolution of the Precambrian mobile bel t of

north Singhbhum. Ind. Jour. Earth Sc i . , CEISM Seminar

Vol., p.205-212.

Rao, N.K. and G.V.U Rao (1983) « Uranium mineralisation in Singh-

bhum Shear Zone, Bihar. I . Ore mineralogy and petro-

graphy. Jour. Geol. Soc. India, V.24, p.437-453.

Sarkar, S.N. and Saha, A.K. (1962) t A revision of the Precambrian

stratigraphy and tectonics of" Singhbhum and adjacent

regions. Guart. Jour. Geol. Min. Met. Soc. of India*

V. 34, p.97-136.

Sarkar, S.C. (1984) : Geology and Ore mineralisation of the Singhbhum

Copper-Uranium belt . Eastern India, Jadavpur University,

Calcutta, 263 p.

- 10 -

GEOLOGICAL MAP OF THE PART OF SNGH6HUM SHEAR ZONESHOWING- URANIUM DEPOSITS

DISTT. SWGHBHUM

PIG 1

- 1 1 -

OEOLOOICAL MAP WITH BOREHOLE LOCATIONS , MOHULDIH .

DISTRICT - SINOHBHUM , BIHAR

SCALEM> O SO CO

I U C K

CHLOWTt SCHIST.

lmm\*\ CHLMITC SCMCIU SCHIST

I V . - . I WWOTC 9CMST WltH TOUMMUNS

I . V . M •WOO MMWTITI OUMTtnt

[•».».«! SOM OWUMC

K0OIN0

5CHUTOV1Y

LOCATION Of •OftCHOLfl

; MOMUUJIHCAM*

F I G 2

- 12 -

BOREHOLE LOCATION PLAN OF NANDUP-TURAMDIH AREA

x \ \C ,J

AJLFIG 3

- 13 -

150*-

100 •so -

0-0 •

50

b l l

1 I 1

f —

7 6A 6 1 t 200

|> r - ; . | 3E/tICITJt 0UAJIT2 SCHIST f- 1 URANIUM Oft*

prrren CKUXIITC QUARTZ SCHIST/FSLU6PATHIC SCHIST

B

Pis. **>

Fit.

Vertical FroJ*«tloa of OreSoutk aloBg AA*.

»t

LoasituAlMl Vertical i ro jcc t lo* of Ore. ko*j aitTuraailk South aloai ! * • .(r«ftr*a«« ! ! • • *r»w« la

- H -

FIG 5b

- --,| Sariaita Quartz Sakiat

i..' I Quartzita a*a Serlaita SakiatK;:-»'*fl (tal«o««) witk touraallma

T H Cklorlt* Quartz Saaist

Uramlua Ora-

It .25 22 2f

. ISO H

FIG saStrlka Saatlbm aloaj tka aorakolas of III Sariesat Mokulalk.

Fie* Sat Strlka Saatloa aloac tka korekolaa of IX.S«ritaat Mokulalk. ( rafaraaaa llaa aratm lm Fie? )•

FIG 6

?2 fo l is _o» tke «ore of the borekoles of Mohuldlk.

- 16 -

FIG;^^.^-j Seri«ite Qttartz ScJ»i»t

I.1. .• J Suartzite a*4 Serivlte S«alst1 " ' n (tal«cs€) wita tour«alia«

^ j Chlorite Quartz S

Uraalua Ore

A Typical Dip Se«tiom aloag tke ¥oreliol«i of Kohuliik.( reference liae Arawa i» Pit.2 )

- 17 -

L-NANDUP TUCAMOlH SOUTH

2>.

TURAMOIH NORTH

m197

Scrl«lte Quartz S«kist |^. _j Ursmlua Ore

Chlorite Quartz Stkist/FeH»p*tki» Stklst

FI68

Dip Sevtlos passlMff tkroujk korekoles of NaiatLup,Tura»4ik Soutk a»4 Turandlh North.

C referem»« l i a ^ l r i m lm Pl«.3 )

6 01J

V«0NUU otrvenn tv «o jtmi

TTHWOOI jo iaviM»

oeit1*

' HVHM' IWWH9WS ISM

'V3UV HKrWHON *m W01N03 TWUMUUS

- 81 -

- 19 -

PROSPECTING FOR URANIUM IN CARBONATE ROCKS OF THE VEMPALLE

FORMATION, CUDDAPAH BASIN, ANDHRA PRADESH

M. VASUDEVA RAO J . C . NAGABHUSHANA A . V . JEYAGOPAL

a n d M. THIMMAIAH

Southern Region,Regional Centre for Exploration and Research,

Atomic Minerals Division,Department of Atomic Energy,

BANGALORE - 7 2

Detailed exploration of the carbonate rocks of the VempalleFormation of the Proterozoic Qiddapah Supergroup has led to theidenti f icat ion of a promising stratabound uranium horizonhaving correlatable mineralization of good width, grade, andextent in 18 l o c a l i t i e s over a stretch of 62 km. Sub-surfaceexploration at two l o c a l i t i e s (Thummalapalle and Gadankipalli)has resulted in delineation of ore bodies with good grade*sizeable tonnage, and thickness down to shallow depths of about150 m.

The exploration methodology adopted, various Integratedtechniques used and guides recognized during exploration*together with results obtained are discussed. Suggestions fordeveloping exploration programmes in similar l i thostratigraphicsett ings elsewhere in the country are also made*

- 20 -

INTRODUCTION

The middle t o l a t e Proterozoic Cuddapah basin has been afavouri te ground for exp lorat ion g e o l o g i s t s and mineral prospec-tors s i n c e as ear ly as 1625, due t o i t s as soc ia ted diamond occure-nces as well as a s b e s t o s , b a r y t e s , and base metal minera l i sa t ion .This bas in has been radiometr ica l ly surveyed s ince mid s i x t i e sbecause many favourable c r i t e r i a for uranium concentration l i k ethe s t r a t i g r a p h l c s e t t i n g c o n s i s t i n g of middle Proterozoic psammo-pel i t i c sediments and chemical precipitates, very fertile grani-toids in the vicinity, and the repeated phasas of igneous acti' ityof both basic and acidic nature, are present. During these earliersurveys, the basal Gulcheru conglomerates resting unconformablyover the Archean gneisses/granites were found to be radioactivemainly due to thorium. During the mid-eighties, samples of phos-phorites associated with the Vempalle limestone, being investi-gated then by the Geological Survey of India, were found to con-tain appreciable uranium. Detailed investigations by the AtomicMinerals Division have brought to light a unique type of strata-bound U-mlneralisatlon in association with the Vempalle carbonaterock belonging to the Middle Proterozoic Papaghnl Group of theCuddapah Supergroup* The mineralised carbonate rock is admixed,at many places, with phosphatic and siliceous material, and hasbeen traced over a stretch of 62 km, wherein about 18 interestingzones are delineated by ground radiometric surveys (Fig.l) .

After the discovery of this mineralisation, a systematicexploration methodology has been adopted, taking Into considera-tion both the field guides and genetic models. Photogeology,airborne gamma ray spectrometrlc techniques and hydrogeochemicalsurveys were adopted during the early stages to cover largerareas in shorter time and to delineate favourable areas fordetailed follow-up investigations. Results of the hydrogeo-chemical surveys carried out In the Vempalle Carbonate rock andthe overlying Upper Cuddapah sediments are discussed in a separatepaper presented in this Symposium. In the anomalous zone*.

- 2 1 -

detailed ground checking by radiometric methods was followedby shielded probe logging of the outcrop areas. Encouraged bythe good strike extensions, width* and favourable analyticalresults, shallow down the hole drilling was initiated in thelater stages. After the Initial success, core drilling wasintroduced to study the subsurface behaviour of uranium minera-lisation.

A detailed account of these different phases of explora-tion that brought to light a sizeable uranium deposit ofencouraging grade and thickness, together with i t s geological setup are dealt with in this paper.

GEOLOGICAL SET-UP

Different aspects of the Cuddapah basin are described In thec l a s s i c work of King (1872) . MLth an object t o provide • newoutlook i n understanding the evolut ion of the Cuddapah basin, arev i sed l l t h o s t r a t i g r a p h l c c l a s s i f i c a t i o n has bean proposedrecent ly by Nagaraja Rao e t a l . , (1987) taking i n t o considerationthe stratigraphy, s tructure and evolut ion of the bas in .

. This s t ra t lgraphlc success ion and the uraniferous horizonsi d e n t i f i e d in the Lower Cuddapah sediments are given below*

Age Group Formation Rock types

? Gandikota cju*rtrite CuartriteTadapatri shale ShaleRil ivendla quartz i te Conglomerate U-minera-

Y Ouartzite llsationLi

msconformity* Wtmpalle limestone/ Stromotolltlc (U-minera-n z o-r>«K«4 shale dolomite, ( l lsation° O % 2 j j l 1 dolomita, mud (CStratabound)2jjxij wrwp stone, chert,s C breccia basicJ s i l l a and dykest Gulcheru quartsita Conglomerate,

Unconformity ^y^^i^tArch- ttMmmmmn*' QCV\LXB/ \ (u-mlncra-aean »asem»nc Gneisses I lisation

\ (fractureI and shearI controlled)

- 22 -

The Vempalle Formation, which i s by far the most importantfrom the point of uranium mineralisation, conformably overliesthe ^ulcheru quartzite, both constituting the Papaghni Gro\(pof the Lower Cuddapahs.

Gulcheru quartzite, the basal member of the Papaghni Groupoverlying the Archaean basement ( i l g . l ) with a profound uncon-formity, consists mainly of conglomerate/grit, arkose and quart-z i te with shale intercalation, and has a thickness of 33 to 280 m.

The Vempalle sediments are mainly calcareous consisting ofstromatolites and dolostone with intercalating quartzites, con-glomerates and chert bands. The estimated thickness i s around1800 - 2100 m (Roy, 1947) • This unit i s traversed by basicdykes. Lower Cuddapah sediments have witnessed magmatic activitymanifested in the form of sub-aerial basic lava flows, s i l l anddyke intrusions.

EXPLORATION METHODOLOGY AND RESULTS

Ground radiometry

I n i t i a l ground radiometrlc checking has revealed thepresence of mineralised carbonate rocks (Vempalle Formation)recording radioactivity of the order of 3 to 10 times the back-ground count and commonly ris ing above 15 times intermittentlyalong a 62 km long b e l t between Komantula in the west andCuddapah in the eas t . Eighteen anamolous zones have been identi-f i e s in th is be l t , with individual outcrops varying from 200 a to1.5 km in strike length and 20 to 25 m in width. The importantloca l i t i e s* where detailed investigations are being carried out,ares Tummalapalle, Gadanklpalle, Rachakuntapalle, and Bakkanna-garipal le . I t has been noticed that high order radioactivityin the carbonate rock i s associated with s i l i ca -r i ch portions,dark bands of s l l t s tones , chert and stromatolites. Radiometrlcassay values of about 200 grab samples from these areas show0.01% to 0.20% eU30Q, with a corresponding 0.01% to 0.22% U30g(«/r)

- 23 -

and negligible thorium. Chemical analyses confirm the radio-metric data.

Shielded probe logging and non-coring dri l l ing

Shielded probe logging of the mineralised outcrops overthe dip slopes indicates average values of the order of 0*02%to 0.03% eu

3O0 over widths of 10 to 25 m. As most of the rockexposed i s along the dip slope and escarpment outcrops are lacking,shallow down the hole (DTH) dri l l ing i s carried out to know thetrue thickness and grade of the mineralised horizon in two promisingareas - Tummalapalle and Gadankipalle - with a dri l l ing intervalof 50 m to 100 m along the strike to intercept the mineralisationat a depth of 10 m to 30 m. With this dri l l ing, a strike lengthupto 1200 m each i s delineated both at Tummalapalle and Gadankipalle.The grade and thickness of the mineralisation varies from 0.02%•U308 x 1.5m to 0.050% « u

3 0 8 x 4.5 nu

Core dr i l l ing

In order to study the subsurface samples with respect tomineralogy, grade and geochemical parameters, core dri l l ing i scarried out In these two areas. The pattern of borehole locationsand results obtained from each of the two areas are given below*

Tummalapalle deposit

In the Tummalapalle area, the uranlferous carbonate rock i ssandwitched between a lower massive limestone (with intercalatoryshale bands) and upper cherty limestone (Fig.2) . The mineralisedcarbonate rock measures upto 20 m, and i s further made up of inter-calatory mudstone, with development of mudcracks. At places*ripple marks and stromatolites are very common in this carbonaterock* A thin lmpersistent layer of conglomerate i s often recog-nised separating the underlying massive limestone and the uraniferoushorizon. The shale unit immediately succeeding the mineralised zoneIs fairly uniform snd typically purple In colour with well developed

- 24 -

partings. Thus, this unit marks the upper marker horizon, while

the conglomerate serves as lower marker horizon for the minera-

lised carbonate rock. All these formations have general east-

west str ike, with low dips of 10° - 15° towards north (Fig.2).

In the f i rs t series, boreholes were drilled at an interval

of 100 ra along the strike to intercept the mineralised horizons

at vertical depths of 50 m to 75 m. This drilling has Indicated

the presence of two bands of mineralisation- the hangwall and the

footwall bands - separated by a zone of lean mineralisation of

3-5 m thickness, and has established the correlatability and the

stratabound nature of the mineralisation over a strike length

of 1.8 km. Encouraged by this, drilling to establish the dip

continuity upto 620 m and to a vertical depth of 150 m has been

taken up at 200 m interval along the strike. Drilling carried

out so far intercepted the mineralised horizon correl a table with

the ooreholes drilled up dip and also along the strike for 1.6 km

(Pigs. 3 and 4). The average grade and thickness of the minera-

lised bands are 0.04IX *V2°8 x 2*2$ m E O r hangwall band and 0*050

x 1.6 m for the footwall band, besides appreciable concentration

of molybdenum (average 300 ppm) in the hangwall band. The bore-

holes drilled in the intermediate scries have confirmed the above

observations*

Gadankipalle deposit

In the Gadankipalle area/ the geological set-up i s very

much similar to that of the Tunmalapalle area, excepting for the

absence of intercalatory conglomerate and poor development of

the hangwall purple shale. The thickness of mineralised carbonate

rock i s 20-30 m. The basic dyke* which i s so prominent at lUmmala-

palle, i s not present in this area. The formations have east-west

strike with low northerly dip of 15°»2O° (Fig.5).

After the initial OTH drilling, which established a strike

correlation of mineralisation upto 1200 m, a block on the western

side of Gadankipalle measuring 500 x 500 m i s selected for core

- 25 -

drilling in 100 m x 100 m grid to know the depth persistanceof mineralisation. Drilling completed so far has thus indicatedcorrelatable ore grade mineralisation, both along the strike anddip.

Three mineralised bands are present in this area* of which

the hangwall band has the characteristics of ore grade minerali-

sation of 0.030% eU308 x 1.5 m to 0.040% eUjOg x 4.5 m (Fig 6) .

Preliminary analytical data on samples from this area have

established the molybdenum content comparable to that at Tummala-

palle.

Drilling in this area i s under progress to establish further

strike and dip continuity of the mineralised zone.

LABORATORY STUDIES ON THE ORE

r

The mineralised carbonate rock comprises alternate bandsof dolomite-rich carbonate and collophane-rich phosphate. Theradioactive minerals - pitchblende, and coffinite - occur eitherwithin the phosphate-rich band or at the junction between thisand the carbonate band. In addition* some suspected organicmaterial in association with pyrite has been identified. TheP-Oc content in the surface samples varies from 5 to 15%, andupto 35% very rarely, whereas in core sample i t i s 1 to 5%.There i s good positive correlation between uranium and p

2°5 i n

core samples, whereas the sane in the surface samples i s insigni-ficant.

Among the trace elements, Ni, Cu, and Mo are present inappreciable concentration as compared to the Clark's values. Ofthese, molybdenum concentration assumes economic significance*

The leachability studies carried out on the surface and

subsurface samples of the mineralised zones by the Mineral Tech-

nology Laboratory, AMD, have indicated leacheability varying

from 60% to 70% and in few cases upto 60% through carbonate route*

Similar studies are also underway at the Uranium Extraction

- 26 -

Division, Bhabha Atomic Research Centre (BARC), Bombay.

Further studies are in progress to achieve improvement

and recovery of associated molybdenum as a bye product.

DISCUSSION

From the data accrued sofar, both on the surface and sub-surface samples, i t has been established that the uranium minera-l i sa t i on , confined to the carbonate rock of the Vempalle Formation,i s str at abound. This l i t h i c unit occupies a dis t inct stratigra-phic position being sandwitched between the massive limestoneand the cherty limestone of the Papaghnl Group. This s t r a t i -graphic control and other associated sedimentary structures l ikeripple marks, mudcracks and stromatolites are important f ie ldguides for locating the uraniferous horizon in the study area.

Exploration by non-core and core dri l l ing methods in theTummalapalle area has resulted in delineating a cor rel a table andcontinuous ore zone of over 1.8 km, thereby establishing substan-t i a l Inferred category uranium ore reserve in the two ore bands.

. In the Gadankipalle area also, same exploration has esta-

blished the continuity of the ore zone* over 1200 m strike length

and 400 m Inclined length along the dip direction. +

By adopting a combination of dri l l ing of non-coring and

coring methods judiciously, the evaluation has been made possible

in shorter time, besides economising the dri l l ing cost to a great

extent.

The correlatabil i ty of ore bands both along the strikeand dip and the high degree of consistency of their grade andthickness are remarkable in the two study areas, further dri l l ingin these two areas i s in progress to establish additional reservesin the inferred category and to convert the rmamrvmrn from inferredto indicated category. The high concentration of molybdenum inthe ore zone (average about 0*03%) i s an additional factor toenhance the economic v iabi l i ty of these deposits. Another very

encouraging aspect is the disequilibrium factor, generally infavour of parent uranium of order of 20-25%, which would enhancethe actual tonnage of the uranium reserves.

As has been mentioned"earlier, the radioactive carbonate

rock has been traced over 62 km strike length and 18 promising

zones identified, with Tummalapalle and Gadanklpalle being the

two, which are under detailed exploration. In the light of expe-

z-ience already gained in the Tummalapalle and Gadankipalle areas,

another five zones which have very good surface indications of

radioactivity, with significant strike length are proposed to be

taken up for further exploration by drilling. These five are

Rachakuntapalle (West), Rachakuntapalle (Bast), Gadankipalle-II,

Bakkaiiagaripalli (B.K.Palli) and Velamvarlpalle. I t i s expected

that the mineralisation in these five zones too would behave

similarly for proving another sizeable deposit of uranium.

It has been seen from the above that exploration by an

integrated approach taking into account the favourabllity criteria

like stratlgraphlc setting, lithology, and structure has helped

In delineating highly promising zones of uranium mineralisation

In the caroonate rocks of middle Proterozoic Vempalle Formation of

the Cuddapah Supergroup.

This unique type of str at abound uranium mineralisation in theVempalle carbonate rock with vast lateral extent and remarkableconsistency in grade and thickness has the potentiality to contributesubstantial reserves to uranlam resources of the country*

There are several mid to late Proterozoic intracratonlc sedi-mentary basins in India* the important among them being the Chattis-garh, Indravathi, Vindhyan, Pakhal and Abujhmar basins, which exhibitsimilar llthostratigraphlc and chronostratlgraphic characters asof the Cuddapah basin. I t i s hoped that a systematic study ofthese basins on the lines carried oat in the Cuddapah basin, wouldbring out many more promising uranium fields in this country*

- 28 -

ACKNOWLEDGEMENTS

The authors are highly grateful to Shri A.C. Saraswat,

Director, Atomic Minerals Division (AMD), Sri S.G.Vasudeva,

Regional Director, Southern Region, AMD, for all the guidance,

encouragement and support extended for carrying out investi-

gation in the Cuddapah basin. They are thankful to S/Shri D, Veera-

bhaskar and K. Ramesh Kumar for discussions and valuable suggestions,

- 29 -

REFERENCES

KING w. (1872): Kadapa and Kurnool Formations,Geol . 3urv. India Mem. 320 p .

NAGARAJA RAO B.K., RAJURKAR S .T . , RAMALINGASWAMY G., and RAVINDRA BABU B. (1987) xStratigraphy, structure and evolution ofCuddapah Basin. Geological Society ofIndia, p. 33-86.

ROY A.K. (1947) t Geology of the Ohone Talukand neighbouring parts, Kurnool d is t r ic t .Geol. Surv. India progress Report (1945-46)(unpublished)

- 30 -

b tOI .OGlCAL MAP OF PARTS OF CUDDAPAH BASIN

SHOWING URANIUM OCCURlINCL'i,

FlG.l

•..V..NHV..L.:C

S.L.T:K

v L.::i:::

+ ... .y v.

-h -I . f -»- I 4-

•RAYACKOTL/ v:

7 1 3 UPANIUM OCCUnHENOE INVCMTALLF. ,UOLOSTWJE 1 PULLIVENIXA QUARrZITE!

FAULT/FRACTURE ZONE

UttttJ KDONDAIR LIMESTONE

jrttl JAMMALAMADUGU ' LIMESTONE

CUMDUM SHALE

I OAineNKONOA QUART ZITE

fi'-VJL'J TAOPATni 3I1ALPW.:~::^l f'ULLIVENOLA/NAGARI QUAMTZITEn = ^ ' MASIC SILLS/VOLCANIC FLOWS

i'•.'•'• IVF.MPALLJ: DOLO5JONC/LIMLSIUNI£/<JIIAI.Erm~r~n cui

Fir,.?

GEOLOGICAL MA? OF TUMMALAPALLE AREACUODArAH. DISTK A.P).

, k*ii: «;jr.

I

CM

- 35 -

TUMMALAPALLE AREA CUDDAPAH. DISTT. (A.P)

•lti.it

Q t

i ^Ml*s^ve umt sroue

GEOLOGICAL MAP OF GADANKI PALLECUC3APAH.DI5TT. (A.P)

AREA

3 13D

11

1 11

_"2 •:' 1 -• *' c-H

P I [ CMERTV LIMtSTONt

, CUDDAPW OlSTT

^ 3.5 »

IT «W.U.i : i ; . . | i ' : i ^ v

EVALUATION OK FAVOURABLE STRUCTURAL FEATURES

FOK URANIUM FROM AIRUOKNE GEOPHYSICAL

SURVEYS OVER PARTS OF MADHYA PRADESH. INDIA

K.I.. TIKU. S.V. KRISHNA RAO and BIPAN BBHARl

Atomic Minerals Division

Department of Atomic Energy

Government of India

Hyderabad - 500 016

The present study focusses on the interpretation of aero-

magnetic and aerial spectrometrir. data of two areas in Madhya

Pradesh, v iz . , 'Bilaspur block' north of the Chhattisgarh basin

and 'Raipur block' situated south of this basin. Both the

blocks comprise different chronostratlgraphic units starting from

Archaean age.

The aeromagnetic map clearly demarcates rocks of

Chhattisgarh Supergroup of Upper Proterozoic age in the Bilaspur

block. Lower Gondwana sediments (Talcher GroupJ occur towards

north and northeast. Deccan traps are exposed in the northwest

in this block. The rest of the area in this block is covered

by Archaean granites and Lower Proterozoic rocks.

The aeromagnetic map of Raipur block delineates Archaean

granite gneisses in the south and the Chhattisjjarh Supergroup

of rocks in the northwest. Some dolerite dikos and Upper

Protorozdir. schists havn also boon dolinoated in tho southwest.

Structurally, two major trends, NW-SE and E-VV have been

reported in the region. The NVV-SE trends represent the foliation

direction parallel to the Mahanadi trend. The E-W structures

correspond to the Satpura strike. Both these structural trends

are identified on the aeromagnetic maps. Four magnetic

linuamonts about 40 km each, trending E-W traverse through

- 57 -

Archaean rocks and Gondwana sediments in the Bilaspur block.A similar lineament is observed in the Raipur block. Aqualitative analysis of these lineaments indicates presence oflinear magnetic sources having mafic to ultramafic compositionat very shallow depths. Many NW-SE faults either terminateor laterally shift these lineaments at several locations. Thus,both the structural trends, i .e . E-W and NW-SE are recognisedon the aeromagnetic maps of both the blocks.

Uranium anomalies from airborne spectrometric data havebeen plotted on the aeromagnetic maps. The distribution ofthese anomalies indicates that uranium mineralisation has apreferential enrichment close to the NW-SE structures, contactzones and near the intersection of E-W and NW-SE structures.It is . therefore, concluded that NW-SE structural features andcontact zones may be promising targets for ground follow-up.

INTRODUCTION

Aeromagneticshas a long history as a method of geophysicalexploration and Is a very Important tool used In any mineralor oil exploration programme. Besides delineating structuralfeatures and lithologlcal units, it plays a significant role inoxploring and identifying potential mineral belts. Recently,Grant (1985) has given the geophysical concept of "OreEnvironments" that can be recognised from airborne magnetometorsurveys due to the characteristic features of magnetic mineralogy.Though, thoro may not be a direct relationship between magnetiteand uranium ore environment, combination of aeromagnetic andaerial spectrometric data may identify potential areas of uraniumore concentration.

The present study deals with the interpretation ofaeromagnetic maps of two areas in Madhya Pradesh, borderingthe Chhattlsgarh Cuddapah Supergroup. Onn area in the 'Bilaspur

- 38 -

Block1 is north of the Chhattisgarh basin and the other

'Raipur Block1 is located south of the basin (Figure 1).

in

REGIONAL GEOLOGY

The two areas under investigation have a similar geological

setting. However. Lower Gondwana sediments occur towards

north and north-east in the Bilaspur Block. Figures 2 and 4

show the general geology of the two areas. The chrono-

stratigraphic relationships (GSI. 1978 and 1979) as recognised

in these areas is tabulated below : -

PERIOD GROUP GENERAL LITHOLOGY

Recent to Subrecent Soil fi laterite withbauxite

Upper Cretaceous Deccan Traps Fissure lava flows

Upper Carboniferous Talcher Groupt(Lower Gondwanas)

Boulder bed.conglomerates,needle shales 8sandstones

Upper Protorozoic

(ChhattisgarhCuddapahSuporgrnup)

Raipur Group Limestones fi shales

Chandrapur Group Sandstones

Lower Proterozoic Granites, doloritedikes, schists

Archaean Grant to - gnolsses,schists, amphlbolltes

M 4 ? »I. 4 4 .4 *

Arttt «•»•

L E G E N D

L«Mtt«« Cn.Cl9T0CCHC>

( T ) DltcM Trty CCMCTACCOUS-COCCHC)

( T ) l i m i t • CrMp (UCUCCOUSI

( T ) U « t r *MrfwM«i (UfPCR CARtONITCROUS-V~^ 10WC* TNIASSIC )

( T ) CklM«llM«rl> C«««,iK t(UfPER PROTEROZOIC)

(T) U»cU«lil.«TCROZOIC )

Fig.1 Location map of areas flown I—^with regional geology

• I'D- • 4 * • * •

- 40 -

Two main directions of foliations have been reported in

Bilaspur block (Rao. 1981). NW-SE foliation direction corresponding

to the Mahanadi trend, is the earlier one. The later E-W

structures parallel to the Satpura str ike, are superimposed on

the NW-SE trend. Cross-folds represented by NE-SW foliation

direction appears to be the resultant of the above two trends.

In Raipur block some schistose rocks and dolerite dikes

are exposed in the south-western part. The strike of the

schistose rocks and trend of the dikes is NW-SE (Figure 4 ) .

Thus, one of tho major structural trend in this area appears

to be NW-SE.

THE BILASPUR BLOCK

Deccan Traps occur in the north-western part of this block

with Lameta Croup of sediments bordering all along (Figure 2 ) .

The Lower Gondwanas (Talcher Group) in the north and north-east

lie directly over the granite-gneiss. The Lower Proterozoic

rocks that have been tentatively correlated with the Lower

Sausers (Rao. 1981) are exposed north of the Chhattisgarh

Supergroup.

The magnetic contour map:

Deccan traps can be demarcated clearly on the magnetic

map of the Bilaspur Block (Figure 3 ) . by the characteristic

magnetic contour pattern. Here, the magnetic contours show

closely packed small 'highs' and ' lows' , the variation of the

field being between 300 to 700 gammas. The smooth magnetic

Hold in the southern part of the Block distinguishes tho

sedimentary formation of the Chhattisgarh basin. The gradual

docrnase of the field also Indicates southerly slope of the basin.

The prominent features in the aeromagnetic contour map

of the Bilaspur Block (Figure 3) are four magnetic lineaments

- 41 -

. ; % • • ^

© Ml CM >M>

© t~..._*©©e

0 jgw/MMuj* 1

O MM»«UMf< j

FI9.2 GENERAL GEOLOGICAL MAP OF BH.ASPUR aOCK

.- . . > i i . i I «

- 42 -

trending E-W. It can be observed that the lineaments appear

to originate from the Deccan Traps and traverse through both

the Archaean rocks and the Gondwana sediments, covering a length

of about 40 km. At most places these lineaments show a magnetic

'lows' with varying order of total magnetic field between 700

and 200 gammas. Considering the induction in the present day

Earth's magnetic f ield, these lineaments indicate the presence

of sources of linear geometry with moderate dips due south

(Parker Gay, 1963: Reford. 1984). The amplitude of field

intensities and their sharpness suggest that they are basic dikes

with very shallow depth of burial.

A low magnetic field of the order of 150 gammas near

Koshani demarcates the brecclated granite. The 'low* may be

attributed to the depletion of magnetic material from the granite

due to brecciation. The south-western contact of this rock type

appears to be faulted by NW-SE fault.

A number of NW-SE faults also can be observed on the

magnetic maps. They either displace the magnetic lineaments

or abruptly terminate them.

Airborne Spectrometric Data:

Contour maps of total counts, U m . Th___ . K% and ratio

maps of this block do not show any significant features. However.

Uranium values varying between 20 and 40 ppm have been piottod

on the map (Figure 3) . These /.ones occur near thn contacts

of the Archaean granites with the Gondwana sediments nnd clnsn

to the NW-SE faults. v*st and north-west of Dandarbarpall.

Tlui ur.inliim annum I Ins north or Khfiimirhi iiro HIHO noar tho

contact of Archaean with Chhattlsgarh and Lower Proterozolc

rocks. Thn Chhattisgarh Cuddapahs nppoar to hnva faulted

contact near Khamaria.

- 43 -

vW*va ••••• . 4«Fig.3 TOTAL INTENSITY AEROMAGNETIC MAP OF

BILASPUR BLOCK WITH URANIUM ANOMALES.mtmm^nmmmmut

- 44 -

THE RAIPUR BLOCK

In this block the north-western part is occupied by the

Chhattisgarh Cuddapah sediments (Figure 4); and in the rest

of the area granite-gneisses of Archaean age occur. The NW-SE

trending schistose rocks and many dolerite dikes are emplaced

in the southwestern part of the area.

Aeromagnetic map:

The results of the aeromagnetic data of the Raipur block

arc presented in Figure 5. The magnetic field over the

Chhattisgarh basin is showing many closed contours irregularly

distributed. This behaviour of the magnetic Hold may be

attributed to the reported ferruginous nature of the Chhattisgarh

sediments here (Murti, 1987).

A nearly circular magnetic 'high' with field intensity

variation from 1200 gammas too 1800 gammas, occurs north-east

of Dhudhwara, within the basin. The magnitude and limited

aerial extent of this anomaly indicate that the causative body

may be of mafic composition, moderately dipping north and of

shallow depth of burial.

The magnetic field over the Archaean terrain south and

east of the Chhattisgarh basin has irregular pattern, showing

that there are many local lithological variations. The contour

trends Indicate E-W strike of the Archaean rocks. However,

strike changes to NW-SE in south-western part of the map where

the outcropping schistose rocks and dolertie dikes also trend

in this direction.

Two linear magnetic anomalies are observed oast of

Mahasamund and Bhoring. Both the anomalies are due to dolerite

dikes.

A magnetic lineament cm be riomarcateel oxtomllng E-W

right across tho mnp (Klguro !i). In the southern part near the

- 45 -

G««*'al (t«l>|<«l m*^ •• Hwfw« Stock fig 1 Toot mfMtiiy AxamMo*IK mof •<Da*x Stock, will) uranwm anomahf*

- 46 -

villages Birgundi and Pandripani. From the estimate or the

source parameters of this lineament It may be inferred that

the causative sources are of basic composition with northerly

dips. Two NW-SE faults have been Interpreted and shown on

the map near Birgundi and Pandripani.

Uranium anomalies:Peak intensity of uranium values obtained from the

spectrometric data. have been plotted on the magnetic map

(Figure 5). Many of them occur near the NW-SE faults close

to their intersection with the magnetic lineament. A string ofuranium anomalies 4 seen at Akalwara and down south all alongthe contact zone between the Archaean and the ChhattisgarhCuddapah.

DISCUSSION OF RESULTS AND CONCLUSIONSThe aeromagnetic maps of Bllaspur and Raipur blocks have

brought out very important structural features in the region.The E-W magnetic lineaments stand out well and have been

interpreted as due to basic dikes. Many NW-SE faults havebeen deduced from their magnetic signature that is duo eitherto the lateral displacement of "magnetic horizon" or itsdiscontinuity. These faults are parallel to the major regionalMahanadi tectonic trend.

Domzalskl (1966) discusses that the dikes represent moatimportant structural features that can be related to the majordirections or fracturing. P. Gay (1972) also observed that thenornmngnotlc llnonnmnts c:nn bo correlated with major tectonicevents. Thus, the E-W magnetic lineaments and dikes In Bllaspurand Raipur blocks may correspond to the Satpura strike in therogion.

From the spectrometric data it is seen that the most

- 47 -

uranium anomalies in both the blocks are located near the NW-SE

faults and contact zones. An important control mjy have been

provided by these faults and have served as channelways for

mineralising solutions. The shearing and fracturing along or

near the contacts between competent Archaean rocks and

incompetent sediments played a role in concentration of uranium.

Thus it is concluded that the Mahanadi tectonic event may

have produced NW-SE fracturing that became the loci for

deposition of mineralisation during the later Satpura tectonic

episode. Hence, the NW-SE faults contact zones and the

intersections of structures in the region appear very Important

locales for ground follow up for further Investigation.

ACKNOWLEDGEMENTS

The authors are thankful to Shri A.C. Saraswat. Director,

Atomic Minerals Division for the encouragement and for permission

to present this paper. S/Shri N.C. Slnha and T. S reed ha ran

have been helpful in preparing the diagrams and typing the

manuscript.

REFERENCES

Domzalski. W.. 1966: Importance of aeromagnetics in evaluation

of structural control of mineralisation: Geoph. Prosp.

v 14. pp. 273-291.

Grant. F . S . . 1905: Aeromagnetics. geology and ore environments:

Geoexpl. v 23. pp . 335-362.

Geological Survey of India. 1978: Quadrangle Maps.

Geological Survey of India. 1979: Quadrangle Maps.

Geological Survey of India, 1962: Geological map of India.

Monkol. M. and Guzman. M., 1977: Magnetic foaturo of fracture

zonos: Cnooxpl. v 15, pp. 173-181.

- 48 -

Murti. K.S.. 1987: Stratigraphy and sedimentation in Chhattisgarh

basin, in "Purana Basins of Peninsular India": Memoi^.

Pub. Geoi. Soc. India. . Bangalore.

Paterson. N.R. and Reeves. C.V., 1985: Applications of gravity

and magnetic surveys: The State-of-the-Art in 1985: Geoph.

v 50. pp. 2558-2594.

Parker Gay. S. , 1963: Standard curves for interpretation of

magnetic anomalies over long tabular bodies: Geoph. v

28. pp. 161-200.

Parker Gay. S. . 1972: Aeromagnetic lineaments and their

significance to geology: American Stereo Map Co.. Salt

Lake City. Utah. USA.

Ran. T . M . . 1981: Structural importance of the rock units seen

in parts of Bilaspur and Khatgora Taluks. Bilaspur district.

Madhya Pradesh: Special Pub. No. 3. GSI pp. 77-79.

Reford. M.S.. 1964: Magnetic anomalies over thin sheets: Geoph.

v 29. pp. 532-536.

- 49 -

INTEGRATED GEOPHYSICAL INVESTIGATIONS FOK UKAN1UM

- A CASE STUDY FROM JAMIRI.

WEST KAMENG DISTRICT. AKUNACHAL PRADESH

R.Srinivas, J.K.Dash, S.Scthuram

K.L.T iku and Dipan DehariAtomic Minerals Div is ion. Departmenl of Atomic Energy.

Begumpet, Hyderabad-500 016

An integrated geophysical approach was attempted for

uranium explorat ion in Jamiri area. Arunachal Pradesh, using

the techniques of magnetic, self-potential (SP) and res i s t i v i t y

p ro f i l i ng , coupled wi th sol id state nuclear track detection

(SSNTOJ, to (Icliiierite favourable structures control l ing uranium

mineralisation in p h y l l i t i c auartzites and quartzites of the

Precambrian Dalinp fnrnii it ion.

Three promising zones of uranium mineralisation were

recognised based on integrated results from these surveys.

Magnetic survey ident i f ied l l thologic contacts and faults in

the area. A high-order SI' anomaly of -900 mV was observed

near the contact of phy l l i t cs in the east and p h y l l i t i c quartzi les

in the west. A very low res i s t i v i t y of 1.0 ohm m and' high

SSNTD values of 120 tracks/nun2 over a background of 20 to

30 tracks/mm2 were also recorded near this contact. These

anomalies are character ist ic of a fault that channelises radon

and gives low res i s t i v i t i es . The SP anomaly may indicate

sulphide mineralisntion and hence uranium mineralisation in

this contact zone nidy be associated wi th sulphides.

The phy l l i t i c quartzitcs wci»t of th is contact

are characterised by magnetic 'h ighs ' ranging from 540 to

- 50 -

900 gammas. Here, SP anomalies are small closures of

-80 to -100 inV. The SSNTD values range between 100 and

120 tracks/mm2. This rock unit (phylllllc qu;irt/.111>)

appears Co host uranium mlner;i I isat ion along with

sulphides at some places where radon anomalies are

high.

A fault in the western portion of the area inter-

preted from the magnetic map separates phyllitic

quartzites in the east and quartzites to its west. The

faulted contact is characterised by a high SP gradient

and SSNTD anomalies of 100 to 140 tracks/mm2. This

contact may also be promising for uranium mineralisa-

tion at depth.

INTRODUCTION

In any mineral exploration programme, an integrated

approach consisting of geological, geophysical and

geochemical methods is usually followed. RadiomeCric

measurements have been -widely applied all over the

world both from air as well as on ground to locate

horizons favourable for uraniun mineralisation, in

addition to their application in prospecting for oil.

and solving some geological problems. The data

obtained can directly lead to in identifying surface

radioactive deposits (Darnley 1981; Killeen 1983 and

Bristow 1983). However, it is not possible to detect'

subsurface deposits using radionetric measurements.

In such cases, non-radlometrlc geophysical methods hive

generally been employed (Darnley 1988; Catzweiler «t al

1981). These methods have been successful in

recognising and identifying subsurface structures and

horizons having physical properties that

may be associated with the uranium mineralisation. In

- 51 -

addition, radon emanomctry as a prospecting tool for

locating subsurface uranium deposits is now a well

established method (Bowie and Cameron 1976) and has

been successful in locating uranium deposits 100 m

below the surface (Gingrich and Fisher 1976).

Modern advances in geophysical methods have made

it possible to explore the geological problems vjith

increased chance of success. The present work is an

attempt to study the applicability of geophysical

methods comprising magnetic, self-potential and

resistivity in association with radon emanomotry, to

discern structures and favourable locales for uranium

mineralisation in Jamirl area, West_ Kameng district,

Arunachal Pradesh. Here, the mineralisation is reported

to occur in the phyllitic quartzites and quartzites of

Dal ing formation belonging to Precambrian age.

The data acquired by these methods have been

processed after applying necessary corrections, contour

•aps prepared and plausible Interpretation offered.

Some structural features with which mineralisation in

the area appears Co be associated, have been

identified.

GEOLOGY

The rock formations in the area are equivalent to

Buxas of Precambrian age. Locally, the lithological

units encountered in Jamirl are quartzites, phyllitic

quartzites, chloritic phyllites and phyllites (Fig.l).

The rock units strike NE-SW and dip 40' to 60' due^ NW.

Uranium mineralisation occurs mainly in phyllitic

quartzites and quart/ites and seems to be structurally

- 52 -

P»1 IMIW MIIKt

f X ) •MMIIK «XHMiKtf

Figure 1 Geological impoccurrences.

of Che area wiCh uranium

conCrolled. In addition, sulphide mineralisaCion

(pyrlCe and chalcopyrlce) is observed wiChin Che

phyllites. Radiomecric analysis of samples shows Che

area Co be predominantly uraniferous.

GEOPHYSICAL SURVEYS

An area of about one square kilometre has been

surveyed by detailed magnetic, self-potential and SSNTD

techniques along the Tenga river valley besides

electrical resistivity profiling over a few selected

t raverses.

Depending upon the accessibility of Che terrain,

traverse Interval o£ 100 m with a station-spacing of 20

m had been chosen for both magnetic and self-potential

surveys, while an interval of 25 m was maintained for

SSNTD surveys. However, closer observations at 10 m and

5 m intervals ^uere recorded near the radioactive

outcrops A, B, C and D in addition to some other

locations wherever it was necessary (Fig.2).

Figure 2 : Geophysical layout map of the study area.

The total Magnetic field values were recorded

using a portable pjroton precession magnetometer. Base

station monitoring was done by another proton

precession Magnetometer at a regular interval of five

minutes. The data corrected for diurnal variations was

reduced to an arbitrary datum level of 47,000 gammas

and presented in the form of a contour map (Fig.3).

Magnetic susceptibility measurements of rock

samples were made both in situ and in laboratory. The

results indicate a low order of susceptibility for most

of the s, iiipl us. However, samples of phvl1Ites and

- 54 -

p h y l l l l L c cpjii rl/.lies rug I sic red u h i g h e r order oi

susceptibility (10 to 20 x 10~6 cgs units) than

quartzites (10 x 10 cgs units) including the samples

j > 1 l l u - d l l | > I I ' I I I I I . i n i i i i i . i I n i l r . r . ' i t l I I I ; I I * I I v i - t i n t « • 1 i > | > : ; .

The same obsciv.il Ion siJlIons ol uugnclic were

used for measurement oC self-potential; the readings so

recorded (in millivolts) were reduced to a common base

and have been presented In the form of n contour map

(Fig.5).

Electrical resistivity profiling was carried out

on a few selected traverses using Schlunberger

electrode configuration with, (a) current electrode

separation of 110 m and potential electrode spacing

separation of 10 m (50-10-50 n) and (b) current

electrode separation of 220 m and potential electrode

separation of 20 m (100-20-100 n).

For SSNTD survey, the area was grldded separately

and auger holes of 50 cm diameter and 100 cm deep were

made at each location. Plastic tumblers with alpha

sensitive SSNTD (Kodak Pathe LR-115, type II) filns

were Implanted in the auger holes and exposed to soil

gas for a period of 21 days during which time Che

seasonal and metereologicol variations are averaged out

(Ghosh and Soundararajan 1984). Thus, the long term

exposure of these films provides an integrated radon

signal. The films were retrieved and processed for

determination of the concentration of alpha tracks. The

track density values are presented in the form of a

contour map (Fig.8).

- 55 -

RESULTS AND DISCUSSION

Magnetic survey :

The uwynelic contour map (t'ijj.3) reveaJs a gentle

field variation in the eastern part and high frequency

nnoni.il Its tow.irds I hi- w r s l c r u sltlc. In ^LMIL 1 i"d 1 . L liu

m a g n e t i c s t r i k e a p p c a r b l o c o i n c i d e w i t h t h e r e g i o n a l

geological strike of the rock formations, which is

NE-SW in the area. A good number of low order anomaly-

closures (20 to 40 gammas) is observed in the eastern

side (east of zero traverse) which may be attributed to

the local variations of magnetic minerals in the

country rock. 'The steep gradients and comparatively

higher magnetic closures in the western portion (west

of zero traverse) may probably be attributed to a

lithological change. Surface geology has revealed the

ZONEIv

Contour Inltrvat JO and WO gommo*AftBltRARY OATUM LEVEL1 '7.000 GAMMAS

Figure 3 : Magnetic anomaly (total field) contour mapshowing anomalous magnetic zone, probablefaults and uranium occurrences.

- 56 -

quartzites Co bo prominent in the west and phyllites in

the east. A NW-SE trending fault marked (Fj-F^ from

the contour pattern of magnetic signature delimits the

"highs" and "lows", thereby indlcal iny .1 I ithoJ ogleaJ

change. The magnetic "highs" observed may be attributed

to the presence of magnetic minerals. These are

correctable with the self-potential and resistivity

responses to be discussed later. A north-south fault

(F--F2) in the western extremity of the area separates

the quartzites in the west from phyllitic quartizites

in the east.

The radioactive outcrops A, B, C and D fall in

the NE-SW trending anomalous magnetic zone (zone I).

This zone appears to continue in the north-east

direction (shown as zone II) with a lateral shift

towards south-east which may be because of the faulting

marked F^-F,. The magnetic anomaly observed between

traverses W2 to 0 and stations 0 to S10, signifies the

presence of a localised body formed due to accumulation

of magnetic material in course of time.

The magnetic susceptibility of rock samples

analysed (both in situ and in the laboratory) shows an

order of 10 to 20 x 10"6 cgs units for phyllites and

phyllitic quartzites and an order less than 10 x 10~6cgs

units for quartzites indicating the absence of any

appreciable susceptibility contrast among the rock

units which could otherwise explain the magnetic'

anomalies of 200 to 300 gammas. This could be because

of the weathered nature of the surface samples studied.

In such a case, there would be a relative concentration

of magnetic material at depth which would produce

anomalies of the above order. For this purpose,

downward continuation (Roy 1966) of a profile AA' was

- 57 -

Figure 4 :

Downward continua-tion of magneticprofiJe (AA1)

attempted which has revealed the depth to the causative

source to be around 25 m (Fig.4).

Self-potential survey :

In the SP contour map (Fig.5) two prominent

anomalies appear: one in the east of the order -900 mV

between W3 and 0. The location of +140 mV anomaly

adjoining -900 nV'anomaly indicates the possibility of

a faulted contact marked as F ^ which has been

clearly brought out in the magnetic contour map. Except

for these two anomalies, rest of the area in the

eastern part shows gentle gradient while high gradient

occurs in the western p a r C indicating a total change in

the ionic concentration from east to west. A similar

change in the gradient has been observed in the

magnetic contour map. A north-south fault is «,«rked

between W9 ond W10 traverses from the contour patterns

- 50 -

Figure 5 : Self-potential contour map of the area.

Profile

A—j a-.,

_ , . - J ? IProfit* Qu I

Figure 6 : Downward cont lnuat Ion oC Sl» profiles (HP'and QQ1)

- 59 -

of SP. This fault clearly demarcates high gradient SPanomalies to its west. Thus, SP map also responds well

in identifying the faults F1~F1 and F2~F2. However, no

correlation between outcropping radioactive occurrences

and SP anomalies could be obtained, unlike the one

observed in the magnetic map. Downward continuation of

two SP profiles PP' and QQ* attempted, has given the

source depth to be of 12 m and 16 m (Fig.6).

Resistivity profilling :

Resistivity profiling, using two separations of

Schlumberger array 50-10-50 m and 100-20-100 m, was

carried out on some profiles In order to study the

lateral variations in resistivity and also to have an

idea of other structural features, if any (Fig.7).

0 Iravcrst

v ^ — - HESISIIVITT PROFILE (S0-1O-SO)

. . . , ' — » HESISTIVItV PPOflLCHOO-70-WI

^ , / ~ \ SElF POTENTIAL PROFILE

Figure 7 : Variation of apparent resistivity ;md SPalong traverses KA and RB.

- 60 -

These profiles show a low order of resistivity in both

the separations between stations Ii5 and W2 where the

apparent resistivity has gone as low as 1.0 ohm m. The

apparent resistivity in general varies between 100 to

150 ohm m in the western portion where the rock

formations encountered are more compact phyllitic

quartzites. In the eastern portion, the order of

resistivity varies from 75 to 100 ohm m in the

phyllites. The low order of resistivity observed

between E5 and W2 may be because of the following

reasons :

(a) presence of a conducting material which gives

an SP anomaly of -900 mV and falls within

this zone (Fig.7) and

(b) the probable indication of a NW-SE trending

fault demarcated from the magnetic map in the

vicinity of this zone.

Thus, the resistivity profiles separate the phyllites

and phyllitic quartzites with the conducting zone In

between them. The conducting zone is the faulted

contact interpreted from the Magnetic nap. No

significant variations in resistivity could be observed

over the known radioactive outcrops.

SSNTD surveys :

The results of SSNTO surveys are presented in the

form of a contour map with contour interval of 10

tracks/mm^ (Fig.8). It is observed that the order of

track density varies between 20 and 50 tracks/mm^ ,

east of the fault F^-Fj. Here, the rock unit

encountered is predominantly phyllite wherein no radon

anomaly is observed.

- 61 -

SCAIE

Contour interval »0 and 20 trodw/mm»r

• i '

Figure 8 : SSNTD contour map .showing .•nom.-.l ous melon

The radioactive outcrops B, C and D of phylliticquartzites occuring within the zone L.are characterisedby track density ano»aly closures of 60 to 200tracks/.*2. This zone falls in between tTie faults F1-Fin the east and F2-F2 in the west. The track densityanomalies nearby F1-F1 range between 60 and X20tr«cks/mm2 «,rked as zone M which registers an SPanomaly of -900 -V with low resistivity of 1.0 oh* «.Similarly the track density anomalies near the faultF2-F2 narked as zone N, are of the order 80 to 140tracks/mn2 and colncide wttn nlgh sp gradlent UranluiJ

mineralisation nay be associated with phylliticquartzites in zone L, whereas the faults Fl-F1 andF2-F2 may be acting as conduits for the radon migrationfrom depth.

- 62 -

Although Arunachal Pradesh provides a heavysurface leaching condition due to continuous rainfall

over 'ong periods, low level of near-surface uranium

and hence low background radon levels are expected in

soil gas. It has been reported (Santos and Gingrich

1983) that these highly leached areas may therefore

show stronger radon anomalies than other areas where

there is more uranium concentration in the soils. It

may therefore be possible to detect deeper sources of

significant uranium mineralisation in such

environments. The zones of high radon anomalies

identified therefore, seem to be promising locales for

mineralisation at Jamiri area.

CONCLUSIONS

Geophysical surveys in Jamiri area have helped in

identifying some favourable structures and zones for

uranium exploration. The Magnetic Method delineated

faults and lithological contacts. Three zones of

proMising uraniuM Mineralisation are demarcated on the

maps. The self-potential Method indicates higher order

of anomaly west of fault F^-F^ where radon values (zone

M) are about five to six times higher than the

background. The high gradient on SP map west of faultF2"^2 i s a l a o favourable because here also radon

anomalies (zone N) are of higher order aligned close to

this fault. The faults way be channel ways for movement

of radon from uranium mineralisation at a depth.

The zone between the faults F\~pi a n d F2~F2

indicates high magnetic anomalies. In this zone of

phyllitic quartzltes small closures of high radon (zone

L) anomalies ranging from three to ten times the back-

- 63 -

g r o u n d v.iluc, L n d Lc;il <_• L h u l t h e |j|iy I I i L I u q u a r t zlt os>

are associated with uranium mineralisation.

ACKNOWLEDGEMENTS

The authors are grateful to Shri A.C.Saraswat,

Director, Atonic Minerals Division, for according

permission to publish this paper. They are thankful to

S/Shri B.M.Swarnkar, P.C.Taneja and Dr.M.A.All for

cooperation in field operations and Dr.P.C.Ghosh for

useful discussions. The services rendered by the

Cartography and Photography section are acknowledged.

REFERENCES

Bowie, B.H.U. and Cameron, J., 1976 : Existing and new

techniques in uranium exploration : in Proc.

I.A.li.A. Synp. on Exploration of Uranium Ore

Deposits : International Atomic Energy Agency,

Vienna, 3-13.

Bristow, Q., 1983 : Airborne gamma-ray spectrometry in

uranium exploration, principles and current

practice : Ind. J. Appl. Radlat. Isot. v 34,

199-229.

Darnley, A.G., 1981 : The relation between uranium

distribution and some major crustal featurea in

Canada, Mineral Mag. v 44, 425-436.

Darnley, A.C. 1988 : The regional geophysics and geo-

chemistry of the Elliot Lake and Athabasca*

uranium areas, Canada : IAEA JC-450.5/3.

Recognition of Uranium provlces. Proceedings of a

Technical Committee Meeting, London, 131-156.

- 64 -

Gatzweiler, R., Schmeling, B. and Tan, B., 1981 :

Exploration of the Key Lake uranium deposits,

Saskatchewan, Canada : IAEA-AC-250/5, Uranium

Exploration Case Histories, Proceedings of an

Advisory Group Meeting, Vienna, 195-220.

Gingrich, J.PJ. and Fisher, J.C., 1976 : Uranium explo-

ration using the track etch method : IAEA/SM-280-

19, 213-227.

Ghosh, F.C. and Soundararajan, M. , 1984 : A technique

for discrimination of radon ( Rn) and thoron

(220Rn) in soil gas using Solid State Nuclear

Track Detectors : Nuclear Tracks, v 9, 23-27.

Killeen, P.G., 1983 : Borehole logging for uranium by

measurement of natural gamma-radiation : Ind. J.

AppJ. Radiat. Isot. v 34, 231-260.

Roy, A., 1966 : The method of continuation in mining

geophysic.il Interpretation : Gcocxplorat ion. v 5,

65-83.

Santos, Jr., G. mid Gingrich, J.K., 1983 : Uranium

exploration in tropical environments using the

track etch system, in Uranium exploration in wet

tropical environments : IAEA. Proc. Advisory

Group Meeting, Vienna. November 1981, 57-72.

- 65 -

CAPTION TOR ILLUSTRATIONS

Figure 1 : Geological nap of the area with uranium

occurances.

Figure 2 : Geophysical layout map of the study area.

Figure 3 : Magnetic anomaly (total field) contour nap

showing anomalous magnetic zone, probable

faults and uranium occurrences.

Figure 4 : Downward continuation of magnetic profile

(AA1).

Figure 5 : Self-potential contour map of the area.

Figure 6 : Downward continuation of SP profiles (PP'

and QQ').

Figure 7 : Variation of apparent resistivity and SP

along traverses KA and KB.

Figure 8 : SSNTD contour map showing anomalous radon

zone*.

- 66 -

N

1 N O E X1-T-r i BUXAS (OOLOMITIC UKSBNCfiRAPHITICr T ~ n SLATES CALCAREOUS OUARIZ1TES)

[?x'?x| 0ALIN6S GNEISS TONGUES

OMJNSS PHYLUTE OUARIZ1TE SEQUENCE

| x X * ] OALING GNEISSES

f A | URANIUM OCCURRENCES

|-"-- ' . | TENTATIVE CONTACT.

1-0 Km=1

- 67 -

W-.D W9 Wg W7 WS WS W4 W3v. 11

LEGENORADIOACTIVE OUTCROP.

El E3 E< ES EC E7 EB E9

- 68 -

wii

WlO w t V. b W7 W6 WS

LEGEND

{ A | oyrcRO*»w

Contour interval 20 and 100 gammasARBITRARY DATUM LEV EL = 47,000 GAMMAS

E5 E6 C'7 E8 £9

- 69 -

Profile Mi urn

2 i t 10 UnitlCONTINUED 0EPlH(H)-*>

80

i H» 1 Unil

I H>2U»it>

' H>) Unilt

Unilf• 0

- 70 -

N10

- N 5

W7 WC W$

LEGEN3SAO'.CiCTIVE OUTCROP

W2

C«kt*yr inUrval Ju an 4 100 mVEt £'

k i .

Nl )A1VHONV 1VI1NJ1M J1J I

A1WONV WI1H31CW J131

- 12 -

0 TraverseRA

TroversRB

Lege n dRESISTIVITY PROFILE (50-10-50)

. . , , '" - -» RESISTIVITY PROFILE(100-20-100)

, ^ / ~ " \ SELF POTENTIAL PROFILE

- 74 -

i:.-.Ti;:-fAi. T;i£vr;ci.Unii:-:-cy:c^ ( i ' "(.'HCLII-RCCK AM A X<.'IIJ:;:VJ.II.».I. T L U .

Ill I'll! EX.-_C.'*ATIu\ ;. F iiASD£JTC:NE-rYi;H UIUMUi". iJ.^CiAi i IUA-JICT-: TC ICWKR KAHADEK GANDL-TCfiE C? I-;EGHA. ,iYA

.7. Dhana Saju , H.C. Bhar/^avaf A.J . oelvan^ and U.K. Virnave^1

Atonic Minera ls D i v i s i o n , Department of Atomic Energy,1 Ban : , a lo re -560 072,2Hyderabnd-5OO 016, and 5 Shi l lonE-793 012

- 75 -

NATURAL

IN Tiin ii.vrLOriATlOu wi'1 oANJoLVa.r.-Tlx-±.

, liiLUA

1 p x

R. Dhana Raju, H.C. Bhargava, A.P. Selvam< and S.N.

Atomic Minerals Division, Department of Atonic Energy,13an-3alore-56O 072,2K}-derabad-5C0 016, and 5Shillong-793 012

Natural Thermoluminescence (NTL) study of whole-rock

and its corresponding quartz-predominant bromoform-light

mineral fraction of the Upper Cretaceous Lower Mahadek

sandstone from the three uranium deposit/prospects of

Domiasiatt Gomaghat, and Irdengshalcap of Keghalava in

northeastern India has shown that NTL patterns on whole-

rock sandstone and its quartz-rich mineral fraction are

very much similar, except for a shift in TL glow peak

temperature by about 50°C toward higher side in case of

the former as compared to that of the latter. Further,

NTL glow curve of uraniferoua (with more than 0.01$ U,0Q)

samples is characterised by two glow peaks — one of low

temperature (LT) at 210°+ 10°C for whole-rock and at

180°+ 14°C for quartz-rich bromoform light mineral frac-

tion, and another of"high temperature (HT) at 260°+ 10°C

and 230°*i 10°C, respectively — , whereas that of uranium-

poor (p^m level) samples is marked by the HT peak only.

These observations, together with rapid and easy way of

taking NTL pattern on whole-rock, point to the NTL tech-

nique on whole-rock ua a potential tool in lar^e scale

exploration for sand3tone-typo uranium deposits, espe-

cially for (a)docipherin;3 the ccmcealod mineralized

zones of even low-level radioactivity, since TL beini;

the net effect of lon^ time radiation exposure, and

Cb)t>redictinr the extensions of unknown uraniferous zones.

- 76 -

i'hermolxin-inescence (TL) of ,eolo.-ic .materials has

found wide a;.»plj.c3tion during the last three and half

decades in different branches of rjeolo;^ like stratigra-

phy (Saunders, 1953; .Kirks, 1953; B'nattachavi.-a et al.,

1976; Ilambi nnd Hitra, 1978), sedinentolO;-;, (Jharlet,

1959), niirieraloj^ (teller, 1954; Manconi nnd HcJougal,

1970; Kaul et al., 197^; Sishita et al., 1974; Hukerji

et al., 1931)* seotheriaometrj (Johnson, 1968), geochro-

nologj (G&nguli and Kaul, 1968; McDou^al, 1968; Nambi

et al., 197^; Pintle and Kuntley, 1982), and or© pros-

pecting (Zeschke, 1963» KcDougal, 196G; Levj, 1977»

'/az and Sifontes, 1978; Ilambi et al., 197(3). In India,

as elsewhere, 'nost TL studies carried out so far were on

TL-sensitoive minerals like quartz, calcite, doloaite,

fluorite, zircon, »nd diamonds (Kaul et al., 1972; 3ha-

ttach^r^a et 3.1., 1976; Uawbi et al., 1-.J78; hutterji et

al., 1->£ji)t whereas :£L stud^ of whole-rock h- s rdceived

comparatively lesser attention (oank iran at al., 1980,

1902, 1983; Jadeyivan et al., 1981; Dhana Haju et al.,

1984). Likewise, TL stud} for prospecting of ores,

thouyn started since earlj 1960s (Zeschke, 19£3)» has

not been aerioaslj applied in India, except for two

recent atudies bj Nambi et al.(i978) and Dbana itaju et

•a., (1904).

Application of TL in uraniuj-i r^eoloi^, co-pared to

other branchun, is a relatively recent one, with raoat

previous studies usini; natural TL as a dosimeter to

detect the i^rosence of uranium ininoralization. These

include the studies b> opirakie et ol., (197?) on a

.jouth I'exaa (U.JA) roll fr'int and Dhana iia$\x et al.,

(1934) on the structurally-controlled h^drothermal

tj i.-e of iin -hbhum (India), witii ootii yfcudiea denonatrat-

ir,%/ an incro'.iue in total TL on «•>. ro:«chir»;; mineralization;

Charlet at ol., (1978), who showed tho use of natural TL

-77 -

to detect buried low-level nineralization, which was other-

wise undetectable b,> other radiometric techniques; and

Hoch.-nan and \pma (iy87), who demonstrated progressive

increase in radiation effects on artificial TL of quartz

(induced by Co gamma radiation) toward the Beverley ore

body (South Australia) in Tertiary sandstones.

In the light of above and as a continuation of the

work on natural thermoluminescence (NTL) study of whole-

rock samples in exploration for uranium (demonstrated

previously on low-grade metamorphic rocks froci the Singh-

bhua shear zone by Dhana xiaju et al., 19Q4-), the present

study of NTL on whole-rock Upper Cretaceous Lower Mahadek

sandstone from three uranium deposit/prospects of Heghalaya

and its quartz-predominant broaoform-light mineral fraction

was carried out with the following objectives:

(i)to find out the application of NTL technique to discri-

minate the uraniferous from non-uraniferous sandstone;

(2)for comparison of NTL on whole-rock sandstone and its

quartz-rich bromoform-light mineral fraction; and

(3)if the NTL patterns on both these are very much similar,

then to propose the technique of NTL on whole-rock sand-

stone as a potential tool iu large scale exploration of

sandstone-type uranium deposits.

tttLMiltLE OF TL AktLliU PO IMAtflUK GAGLOGX

The principles of TL as applied to studies on uranium

geology are exhaustively given by Hochman and Ypma (1987),

and here only important points are described.

Thermoluminescence'(TL) is the phenomenon of emission

of lisht fro-n a crystal previously irradiated, either by-

exposure to nMturallj occurring radioactive minerals in the

field (natural TL) or by exposure to artificial radioaotiv*60

sources in the laboratory, like CO gamma rays (artificial

TL). When an ionizing radiation like 3amJ>a raj enters a

crystal, it dislodges electrons from their 3tonic positions

- 78 -

resulting in formation of free electrons and electronic

holes or sites which have lost an electron. Although

raost electrons .'.nd holes recombine immediately, a small

percentage will, however, be trapped on substitution^

and structural defects. Thus in quartz, the most widely

used mineral in TL investigations, these holes may be

trapped on Al^+ sites and electrons on vacant oxygen

sites. These charges, once trapped, can be released

by heating the crystal. Once released, the holes and

electrons will recorobine, which maj produce a pulse of

light when recombination occurs at a colour centre.

Sucn emission of light is measured with a photomultiplier

and recorded as a ^low peak. As release of trapped'

charges occur over a range temperatures, a number of

glow peak3 results and these constitute a glow curve.

The intensity and shape of the TL glow curve depend

on a number of factors like the number and t^pe of

defect centres capable of acting as traps and their

occupancy rate, which is a largely a function of ionizing

radiation. As charge occupancy rate affects the strength

of the TL signal, TL has been used as a dosimeter to gain

meaningful information relating to present uranium posi-

tion. This is the operative principle behind natural TL

measurements used in the studies on uranium depdsits.

Sample Preparation

Each sandstone sample, after waoiling for removal of

any dirt and drying, W03 powdered to -100 to +1HQ mesh

size (A.JTN). Representative portions of this were taken

bj coning and quartering for TL studj of whole-rock as well

aa quartz-predominant bromoforra-liftht mineral fraction.

The mineral separation was carried out using normal proce-

dures like desliming, acid treatment to remove any shell

matter, magnetite removal bj hand magnet and finally

3oparation of lijhfc mineral fraction by bromoform(Sp. Gr. 2.8>.

- 79 -

Instrument Set-up

The instrument set-up and procedure of the measurement

of TL are essentially same as given in Dhana Raju et al.,

(1984-). Thus, the set-up includes an arrangement for heat-

ing the sample with the sample heater made of a non-corrosive

material, Kanthal and a stx'ip of 15 x 10 x 1 mm central

depression for placing the sample, a thermocouple spot-

welded to the beater strip to determine the temperature

profile, a temperature programmer (made by BAflC, Bombay)

for linear heating of the sample strip and an EMI 9514 B

photomultiplier with S-11 characteristics, and a two pen

•Omniscribe' recorder with four selective chart speeds and

five sensitivity ranges for monitoring the photomultiplier

output and the temperature.

A representative portion of 30 mg of each sample was

heated on the heating strip from room temperature to 400°C,

at a uniform rate of 5° s and TL intensity was recorded

in arbitrary units. Necessary precautions were taken to

avoid the effects of light* ultraviolet radiation, and

other sources during sample preparation and thermal read-

out. For all the samples, background (36) curves were

taken as a routine, and.it was found that the level of

BG was negligible compared to the signal, thus ensuing

that thermal radiation did not alter the signal to noise

ratio. Each sample was repeated four times to get the

average temperature and intensity of slow peak. Even

though interference from tribo- and chemo-luminescence

cannot be ruled out completely, as the measurements are

qualitative and studied under identical conditions, the

final conclusions arrived at will hold good.

- 80 -

.-{adiometric Assay nnd Petrography

Each sample was radiometrically assayed by gamma ray

spectroraetry for its eU,0g, ^x^s* and ThOp contents on

about 400 to 500 grn powdered material. Numerous thin and

polished-thin sections of sandstone samples were studied

in both transmitted and incident lisht for their petro-

^raphic and iainerajira;>hic aspects (details given elsewhere

in Dhana ^aju et al-, 1989). Also the bro-i-oforra-light

mineral fractions of the samples were examined under a

binocular stereo microscope for noting the relative

proportion of light minerals like quartz and feldspars.

GEOLOGIC SETTING

The ShilLocg plateau of Neghalaya, bounded in south

tr. the Dawki fsult, in east and northeast by the Haflong

fault and in north and west by the Brahmaputra river, is

separated from Peninsular shield (more precisely from the

Singhbhum craton) by the Garo-gap. The regional strati-

graphic sequence is as follows*

Alluvium

Younger Tertiary J KopillisFormation

Early Tertiary 1 Sheila Formation (alternatingFormation coal-bearing sandstones

and- limestones)

Upper Cretaceous t Langpars (calcareous.sandstone)irA^,^^ Upper Mahedek sandstoneFormation ^ r H a h a d e k s a n d s t o n e

Jadukata conglomerate

Jurassic » Sylhet Trap

Precambrian t ijhillong Group metasediment3Basement granite/gneiss

The regional .jeolocical set-up and distribution of

different rock types tend to surest that marine trans-

creaaion had taken place from south during the Upper

Cretaceous period, which resulted in the deposition of

- 81. -

very thick sequence of (about 200 m) purple coloured Upper

Mahadek sandstone. This is preceded by deposition of the

Lower Mahadek grey sandstone, which is mainly fluviatile

in origin. Uranium occurrences are mostly confined to this

fluvial facies, and include the already established uranium

deposit at Domiasiat and prospects at Gomaghat and Pdeng-

shakap (Pig. 1).

GEO.0OCAI MAP OF PARTS Of KHASI £ JAMTU M I S . MEGHALATASHOWWG RABOACTIVE OCCURRENCES

j k > |,-T^^^Si^^^

- :==fc= ji I- I- I- ^ f l» 1 - F' H ^i N e i x

>^H Mil ' .'

The Lower Kahadek sandstone from l>omiasiat, Gomaghat,

and Pdbngshakap areas of Meghalaja, on wnich the present TL

studj i3 carried out, ia a [,Tej coloured, friable to hi[;hlj

compact, fine to coarse and occasionally very course grained

(pebbly), 'foldspathic/quartz arenite1, with very little

matrix but predominant quartz and minor feldspar claste,

either oet in cement and clay or erain-supported. The cements

include major amount of organic matter, lesser biogenic and

colloidal pyrite, and occasional calcite, silica and gleuco-

nite (only in. Gomaghat arsa), while rartrix includes micas

and chlorite. Accessories include muscovite, almandine-rich

garnet, zircon and raonaaite. Radioactivity of the sandstone

- 82 -

is mostly due«to uranium present inv the form ultrafine gra-

nular pitchblende intimately associated with low rank orga-

nic matter and p;,rite, admixed U in organic matter, and

minor brannerite, coffinite and zircon. Further details

on petrography and mineragraphj of the sandstone are given

elsewhere (Dhana Raju et al., 1989).

RtSJULTS Alii) DISCUbSIGU

Details of the NTL glow peak temperatures of both

whole-rock sandstone and corresponding quartz-predominant

bromoform light mineral fraction along with U,Og ( /Y )

content are given in Table I. As the TL ^low peak intensity

or height is found to be not having any systematic relation-

ship with U,Og content in both whole-rock and quarts-rich

mineral fraction NTL patterns, the same is neither given

nor discussed here.

NTL in relation to radioactivity

An examination of the data in Table I reveals that

uraniferous samples with nor* than 0.0i£ U.Og &/*) *7e

characterized by two TL glow peaka in both patterns of

whole-rock sandstone (210°+ 10°C and 260°+ 10°C) and it3

corresponding quartz-predominant bromoforo-light mineral

fraction (180°+ 14°C and 250°* 10°C). In contrast, the

U-poor samples with 3-4-6 ppia U,0g (sample numbers 6, 11,

12, 13, nnd 1d) are marked bj onlj one TL 5I.0W peak at

higher temperature of 260°+ 10°C for whole-rock and

°£ 1U°o for quartz-predominant bromoforn»-light mineral

fraction, ^s the high temperature (il'S) , low peak is

common to both uraniferous and uranium-poor samples (260°0

for whole-rocx and 23O°C for quartz-rich mineral fraction),

and onlj in case of the uraniferous samples with more than

O.O1;6 U2C0, there is an additional low temperature (LT)

glow peak (210°0 fow whole-rock and 160°C for quarts-rich

fraction), it appears that the LT glow peak of NTL can be

Table I. NTL slow peak temperatures of whole-rock sandstone

and its corresponding quartz-predominant broraoform-

light mineral fraction, together with U,0g content

Sample U,0flWhole-rock Quartz-predominant

mineral fraction

A. Bomiasiat area123456

0.015 *0.039 'f>0.011 *0.058 *0.14- %48 ppra

B. Gomaghat area

78910111213

0.033 *0.061 5*0.037 #1.00 #12 ppa8 ppm38 ppm

C. Pdengsbakap area

1415161718

0.096 #0.005 *0.044 *0.13 %26 pp«

215215210210210

205215200200

200205205205

andandandandand250

andandandand260255255

andandandand265

255270255255270

255260250255

250250255250

185190185185190

185190185168

184185194170

andandandandand235

andandandand225225230

andandandand222

235240240235220

225230225220

220220235220

Mean

Uraniferous earn-Ipies (* level) I

U-poor samples j(ppm level) I

| 210°+ 10260°+ 10

| 260°+

°c *n10°C

180®!• 14°C230°+ 10wC

230°+

and

10°C

Table II. U,0Q, ThOp, and K contents of whole-rock and NTL peaks

SI.No. ThO, K Whole-rock (°C) Quartz-rich portioncm2610111617

0.089*48 ppra1.00 *12 ppm0.044*0.13 *

0.006*37 ppm0.018*19 ppm0.015*0.034*

1.0*1.0*

1.0*3.0*1.8*

215

200

205205

and250and260andand

270

255

255250

190 and 240235

168 and 220225

194 and 235170 and 220

- 84 -

ascribed to irradiation of samples b} exposure to naturally

occurring radiation of uranium. 3uch a aarked presence of

LT glow peak in the NTL of whole-rock sandstone sample can

be taken advantage of. in discriminating uraniferous zones

from the U-poor zones during large scale exploration for

sandstone-type uranium deposits, especially, for (i)deciphe-

ring the concealed mineralized zones of even low-level, as

TL bein£ a net effect of long time radiation exposure, and

(ii)predicting the extensions of unknown uraniferous zones.

Comparison of NTL of whole-rock with that of quartz-rich

mineral fraction

Data on the HTL glow peak of both whole-rock sandstone

and its corresponding quartz-predominant bromoform-light

mineral fraction (Table 1) demonstrate that both these are

verj much similar in having only HT glow peak for U-poor

samples and both LT and HT glow peaks for uraniferous

samples. The only perceptible difference, however, ia a

shift in peak temperature toward higher side by about

30°C in case of NTL on whole rock.* This shift in both

LT and HT glow peaks in case of the NTL on waole-rock

could be the effect of other minerals associated with

dominant quartz like feldspars and Muscovite present

as clasts, as well as cenent and matrix ot the sandstone.

As the WTL patterns on whole-rock and its quartz-rich

mineral fraction are found to be very xuch similar for

both uraniferous and uranium-poor samples, and as the

NTL on mineral involves laborious and time-consuming

reparation, the NTL of whole-rode, which is simple and

rapid, is preferred to that on separated mineral,

especially when a large number of samples need to be

studied during exploration for uranium.

ttolative effects of U, Tb, and K on NTL

In order to evaluate the relative contribution of U,

Th, and K to the observed NTL, the NTL slow peak tempera-

- 85 -

tures of whole rock and corresponding quartz-rich bromoform-

light mineral fraction are compared with the content of

radioeleraents (Table II). Thus, amongst the uraniferous

samples, those with the lowest and highest contents of

Th (sample no. 2 with 0.006* ThC^ and sample no.17 with

0.034# ThOp) and K (sample no. 2 with 1# KpO and sanple

no. 16 with 5# KpO) h-ive practically no difference either

in LT or HT glow peaks of the NTL on both whole-rock and

quartz-rich broraoform-light mineral fraction. On the

other hand, the U-poor samples with ppm level U (sample

nos. 6 and 11) have onlj the HT glow peak, whereas the

uraniferous samples have both L'S and HT glow peaks in

the NTL pattern of both whole-rock and quartz-rich

bromoform-light mineral fraction, indicating that such

variation in TL temperature is mostly due to uranium.

CONCLUSIONS

(i)Natural Thermoluuinesconce (NTL) studj of whole-'

rock sandstone and its corresponding quartz-predominant

broMoform-light mineral fraction from three sandstone-type

uranium deposit/prospects of Dooiasiat, Gomaghat, and

Fdengshakap of Meghalaja in northeastern India has shown

vivy similar TL patterns for both. The onlj difference

is a shift in TL glow peak temperature by about 50°C

towerd higher side in case of whole-roc'< as compared to

that of the mineral fraction.

(2)NTL of uraniferous samples with .)ore than 0.01 5

IUO^ is charicterized by two TL slow peaka — one at

lower and the other at higher temperature — , whereas

thut of U-poor samplea io marked by onlj one TL glow

peak at higher fcemperiture. Thus, the ui'uniferous oamples

h.ave two UTL t;low [p.-acg .-it 210°^ 10°C and 260°+ 10°C in

caae of wnole-fock :;;i!i.i.;t;une and at 18C°+ 14°G ritia °

10 0 for quartz-rLCii iix.jiitl fraction iii contrast to onlj

the hi^h tetiper.ituro P«;IK ° o °

for the U-poor samples.

- 86 -

(5)A co..";.>ari3on of radioactivity in terras of U, Th,

K contents with the observed TL glow peaks points out

that the IJTL is rr.ostlj due to uranium.

(4)As the NTL patterns of whole-rock and of quartz-rich

bromoforra-light mineral fraction are very much similar in

furnishing information reijardinf; the mineralized and barren

nature of sandstone, and as the KTL on whole-rock being

rapid and easv to take without involving either laborious

and tine-consuming mineral separation or costly irradiation

by exposure to artificial radiation source in a laboratory,

it is, therefore, proposed here to use this technique of

NTL on whole-rock as a potential in large scale exploration

for sandstone-type uranium deposits.

(5)Since TL being a net effect of long time radiation

exposure, it is possible to use the technique of NTL on

whole-rock to decipher the concealed mineralized zones of

even low-level, which otherwise undetectable by usual

radioaetric techniques (viz. Charlet et al., 1978)* and

to predict the extensions of unknown uraniferous zones

that might not have been intercepted in long-interval

drilling operations.

We sincerely thank 3hri A.C. Sarasw.-it, Director,

Atomic Minerals Division (AMD) of the Department of Atonic

Energy for his constant encouragement and permission

to present the paper at the national Symposium on 'Uranium

Technology' at the !3habha Atomic Uese:ir3h Centre (BAHC),

3ouibay from 1Jth to 15 th December, 1999* Our thanks are

also due to Dr. 3. Viswanathan, ohri 6.G. Tewari, and

iinri H.M. Vanna of tho AMD for timely support, and to

the organizers of the ojmposiun at BAJ(O for providing us

an opportunity to ^roaent tho paper.

- 87 -

KEFEHENCES

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- 88 -

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- 90 -

HYJROGHOCKEKICAL EXPLORATION FOR URANIUM: A C.-.JS 32UDY

FROM THE CUDOAPAH BASIN, ANDHRA PRADESH

R.P. Singh , P.K. Jain , B.R.M. Kumar1, S.S.Rao ,

A.V. Patwardhan , and S.G. Vasudeva

Atomic Minerals Oivision

Department of Atomic Energy

1 3angnlore-560 072, 2 Kagour-440 OO1 and 3 Hyderabad-500 016

Hydrogeochemical surveys in the southern part of

the miadle Proterozoic Cuddapah basin, comprising the

Cuddapah (mostly arenaceous and argillaceous) and the

Kurnool (mostly calcareous) Supergroups were taken up on

3-Year Project basis. Nearly 2,30O samples collected in

an area of 4,325 aq Jem during the first year of the Project

were analysed for U, conductivity, pH, and various anions,

viz., CO2", HC03", Cl~, SO42", and cations, vis., Ca2+,

Mg2+, Na+, and K+.

The data indicate that groundwaters from the

quartzitic terrain contain low uranium (2.5 to 4 ppb)#

whereas those from shale and limestone terrains contain

comparatively higher uranium particularly the Upper Kurnool

sequence (Koilkuntla Limestone » 15 ppb, Nandyal £hale

= 1 3 ppb). The U/Conductivity ratios for these shale and

limestone unity r<mge front O.OO6 to O.OO7.

ot-iti ticiil -;nd graphical evaluation of dote ao

woll u.i contouriny of ur-.nlum <<n<i U/ConU»octivity values have

helped in delineating .sixteen anomalous zones, narrowing

down the target area from over 4,000 sq km to 169 sq km for

follow-up studies. Most of the anomalous zones, numbering

elevert^confined to the Upper Kurnool Formation, with an

avoracie uranium content of 50 ppb ( n • 158), and the

remaining five zone;; being confined to :;hale units of

Lov/er md Ux por Cuddapah Groups.

- 01 -

INTRODUCTION

Regional geochemical surveys have been found to be

extremely useful in locating many important deeply buried

uranium deposits in the major uranium-producing countries

such as Canada, United States of America and Australia.

The middle flroterozoic Cuddapah basin was chosen

tor iuch surveys based on several favaurability factors

^uch as: (l) the closed nature of the basin (2) a urani-

ferous fertile granitic provenance surrounding the basin

(3) presence of black shales indicating the euxenic

conditions during deposition of lower and upper Cuddapah

sediments (4) intense igneous activity both in the Lower

Cuddapah times as well as post-Cuddapah times and (5) the

presence of a major unconformity at the base of this

sedimentary basin.

Further, the. middle Froterozoic character of the

basin, in which period an important world-wide orgogeny

(the Mu'J oriian orogeny) has played a major role in the

formation of many major uranium deposits of Canada and

elsewhere, together with the structural deformation and

tln-j therrr.cil episode; of the bruin th-'. accompanied the

-ia^tern Ghat oro-jony make the Cuddapah baoin as a prime

turgut for uruni-un exploration. In view of this, regional

geochemical surveys were taken up on a Project basis

in the middle Proterocoic Cudueijah basin, with the main

objective of rapidly evaluating a substantial part of the

basin and to identify ootential uraniferous areas for

follow-up studies. Results of these studies are dealt

with in this paper.

GEOLOGICAL SETTING

(a) Regional geology

The area under study forms the southern part of the

uiid-P roterozoic crescent-shaped Cuddapah basin of Peninsular

India (Fig. l). This basin is 440 Jan* lone; ond has a maximum

v/idth of 145 km in the middle, covering an area of 44,500

sq.kjn The basin contains over 12 sq tan thick sediments and

volcanics. It consists mainly of ortho-quartzite-carbonate

suite intruded by basic to acid volcanics in the lower part

and siliceous shales with quartzites in the upper part*

The western and southern margins of the basin are

marked by a profound unconformity, with lower Cuddapah sedi-

ments resting on the Archaen Peninaul<ir gneiusic complex.

In this basin, sediments of Cuddapah and Kurnool Supergroups

are preserved. The former is predominantly arenaceous

to argillaccou.; wi'ch jubordin..'.te Ciilc ireou- to dolomitid,

while the Kurnool supergroup con.iits of carbonate facie.;

sediments with subordinate fine elastics.

- 93 -

The geological succession of the area (moditied

King* 1872), mostly followed in our work is as follows:

1 t

1

•

1

1

I

1

%

1

Middleto •U.Prote- «rozoic •

(1600-600 m.y) ,

•

i

a

1

i

KURNCOL

SUPSR-

GROUP

i

| KUNDAIRGROUP

PANEKGROUP

1

JAMMALAMADUGUGROUP

GROUPI

1

UNCONFORMITY

•, <\.I«jiNi*

N/-iNDYAL OH/J1.EKOILICUNTLA LIMESTONE

PINNACLED QUARTZITSPLATEAU QUARTZITE

OWK SHALE

N/iRJI LIME SI' OWE

QUARTZITB & C0H5L0-

3RISAILAM QU.iR'i'ZITES

' GUCUt

GROUP

N.vLLA-KALAIGUOUt

VATHIGXCUP

P..PAOHNIGi.OUP

KOLUI^ULA SHALES

IRL/JCONDA w'UAKT

CUMBUM

BAIKEKKOHDA TjU, lUIV.IY

PULL. J1P£T/ T« 4UP. .TKI J

H. .a .RI/PULVL'tlDI^» iUA

AKCHAEN

("T 2600 m.y)

JUL.-.;< CJN.ivji^ o r Gu

3JE3 &

lM^-rOHE &JIL'-LE

GULCHSaU QUAUTJIl'L'S

I T i - /I'i'M

- 94 -

Major igneous suites associated with the Vempalle

and Tadpatri Formations in the western and southern parts

of the Cuddapah basin are dolerite, picrite and gabbro

sills, basaltic flows and tuffs. Kiraberlite dykes and

syenite stocks have been reported in the Cunibuin shales.

Post-Cuddapah intrusives in the Cumbum shales are reported

in the eastern margin (Nagaraja Rao et al,# 1967) .

(b) Local geology

The area under study comprises (a) the Vempalle

Formation of the Upper Papaghni Group (b) Pulvendla/Nagari

quartzite and Tadpatri shale of the Cheyyair Group, and

(c) Bairenkonda quartzite and Cumbum shale of the Nalla-

malai Group, The Tadipatri shale and Bairenkonda quartzite

are both unconformably overlain b*y the Nandyal shale and

Kollkuntla Limestone of the Upper Kurnool.

The Lower Cuddapahs in this region have a general

strike ranging from E-W to NE-SW, while the Upper Cuddapahs

and the overlying Kurnools have a general NNW-ssB strike.

The Kurnool3 have almost flat dips, while the Cuddapahs

have dips ranging from 15°to 45°toward N or NW. The Cumbum

shales in the e.ist exhibit steeper dips and they are

tightly folded.

- 95 -

C.^ SURVEYS

Sample collection

These surveys consist of collection of ground

water from all available wells representing various litho

units in the area, i.e., from the basement granite to

the Xurnools. A total of 2277 well water samples collected

from an area of 4325 sq Ian were chemically analysed in

the mobile geochemical laboratory of the Southern Region,

Atomic Minerals Division(AMD) for U, Conductivity,

Ca2+, Mg2+, Na+, K+, CC^~, HCO~ , Cl~, S0 42" and pH.

Analytical Techniques

Uranium was determined by the Scintrex UA-3

using fluran or sodium hexametaphosphate buffer. Sodium

and potassium were determined by the Elico flame photo-

meter. Calcium and magnesium w*re together determined by

titration against EDTA using Brichrome 3lack T indicator

at a pH of 10. Calcium way then estimated separately

using Patton end Keeder indicator at a pH of 12, and

magnesium waj then obtained by difference. HCO. <nd CO,

were determined titicLmetrically against standard HC1 using

methyl orange :md phenolpthalene indicator by differential

method. Chloriae was al.;o ieterminod. tltrimatrically

against ataudani oilver nitrate using potassium chromat*

as indicator* Sulphate was determined nephlometrically

- 96 -

by precipitating as barium sulphate using barium chloride.

Detection end precision limits for each element/radical

are as follows: 0.05 ppb +1036 for U, 1 ppm,+ 1O.% for Na

dnd K, and 10 ppm + 5% for the rest.

Conductivity, in terms of micro Mhoc/Cm, was deter-

mined by conductivity bridge, imd pH by pH meter.

RESULTS ^IJ DI3CU3.JICN

The chemical a^oay data were class!ficd lithology -

wise and evaluated statistically. Summary of the data

pertaining to all the major ions aa also U, Conductivity

and U/Comluctivity ±a given in Table I. Statistical

evaluation of the data pertaining1 to conductivity and

U/conduct vity is; ihown in Table II. Of different parameters

the rLitio of U/concluc t ivi ty has been particularly chosen

for ev.ilu tion because, conductivity being an electrical

property »irectly related t^ the total Uissolvetf aoilda,

its ratio with uranium can be utilised for normalising

th<- seasonal fluctuations that affect the con<r

contrition ot U -.long with other -.li :-:;olvec: material*

- 97 -

TABLE II, U. Conductivity and U/Conductlyity of well waters from the Cuddapah basin

UNIT

1.Basement •Granite

2.QulcheruOuartzite

3.VempalleLimestone

4.PulvendlaQuartzite

5.TadpatriShale

6.Bairen-kondaQuartzite

7*CumbumShale

S.NandyalShale

n

20

14

113

71

388

9

612

926

, Mean

73

2.5

8.5

4.0

8.5

3.4

9.2

12.6

U ppb

, Std.t Dev.

83.6

1.8

7.6

3.7

19.0

1.8

12

14.4

, Thre-, shold

240

6

23.7

11.4

46.5

7.0

33

41.4

CONDUCTIVITY(micro Mhos/Cm)

, Mean

1695

489

910

650

1500

1014

1201

1829

, Std.t Dev.

880

220

435

260

1232

448

467

1660

O.

0.

0

0

0

0

0

0

U/Conductivity

Mean

044

055

007

.006

.006

.003

»007

.007

, Std. ', Dev.'

0.04O

0.002

0.005

0.004

0.008

0.001

0.006

0.004

Thre-shold

0.124

0.009

0.018

0.014

0.022

0.005

0.012

0.015

9.KoiI-KuntlaLimestone 77 14.7 18.0 50.7 2770 3662 0.006 0.007 0.020

- 98 -

The data given in table II indicate that ground-

waters from Psammatic sediments (quartzites) show low

(2.5 to 4 ppb) uranium content, while those from shale

and limestone terrains contain relatively higher uranium

values, particularly the Upper Kurnools (Kandyal shale •

13 ppb U, Koilkuntla Limestone 15 ppb U in well waters) .

The average U/Conductivity ratiOvranges for these shale

and limestone units from O.OO6 to 0.007.

The well waters from the granite basement have U values

ranging from £1 to 385 ppb, with an average of 84 ppb

(n m 20), which is quite high even for a granitic terrain.

This is particularly so in the well waters near the Raya-

choti village in the southeastern part of the area under

study, and here the gr&nite-Cuddapeh contact needs to be

investigated in more detail. Vtork on this is In progress.

Data plotting

The urunium values of well waters when plotted in

different maps (not shown) tvive indicated 16 anomalous

zones and these are depicted in Figure 2.

A summarised account cf these anomalous zones Is

(jivon in Table III.

- 99 -

TABLE-III. Anomalous zone3 of Uranium deULaaeated in the

Cuddapah basin by

Litho-Unit

Cuddaoahs:

1. TADP&TRISliALZS

2, CUMDUMSHALES

KUHNOOLS

1. NANDYALSHALES

2. KOIL KUHTL,*LIMESTONE

i'Gi:..L

No.ofanomalouszonesdeleneated

1

4

10

1

16

Hydroqeochemical surveys

AreaOCm2)

30

3O

91

18

169

No.ofsamplesanalysed

28

54

126

32

240

MeanU ppbin wellwaters

44

18

.54

35

42

Thud, out of 4325 sq-fefli investigated by hyttro-

geochemical jurvey3 only 169 sq km h&ve been found

to be anomalous. It ia also of interest to note that most

of the inomalie-j lie close to aome major river courses,

which themselves follow jome prominent lineamentJ.

- 100 -

Based on the computed threshold ond background

concentration values for U suitable isochems are constructed

to define the geometry of the anomalies. One such composite

isochem map of uranium and U/Conductivity is depicted in

Figure 3.

Sample distributions are presented through histograms,

v/hich are generally shewed for areas of mineralisation,

and tend to be lognormel for background aress. Percentage

cunrunul i.tive frequency curves are used for the purpose of

finding bnekcjroum". and threshold values, i.e., 50th per-

centils enci 95th pcrcentile values,respectively. Class

interval and number of classes for these purpose are

chosen to incorporate all the information. Ursnium geo-

chemical dota differ from Gaussian distribution due to

heterogeneity and polymodalicy. Various transformations

are applied to thase types of data to approximate them

with a normal distribution amongst which the logarithmic

tronaformation is the most popular and commonly usec.

Some selected histograms and thair percentage cumulative

frequency curves ure shown in Figures 4a and b.

Ko.3t of the anomalous zones are confined to the N.-indyal

slide unc Kurncol Supergroup in terms of number, dimension

and area. Older fertile granites and mineralised Lower

Cuddapah sediments might h^ve acted as provenance for

accumulation and concentration of uranium in the Kurnool

sediments.

- 101-

Parallelism of the trend of the anomalous zone3

with major lineaments conspicuously followed by rivers ana

their proximity with river courses are implicative of

the role of structure in the mineralisation process.

FUVUKE PROGRAMME

Regional hycirogeocheinical surveys in the remaining

area of the basin will be continued. The delineated

anomalous zones will be taken up for further detailed

radon emanometry and Solid State Nuclear Track Detection (J

techniques to further narrow clown the target. In addition,

systematic aoil sampling will be under taken t-.o supplement

the evidence of mineralisation. The generated data will

be statistically treated with the help of available soft-

ware to facilitate the interpretation and better under -

standing of the geochemical model.

CONCLUSION

Hydro<jeochc:niicc»l surveys have resulted in delineating

several anomalous zone3 under thick soil cover, 'founder

lithounit- ot opr> r Cuddapiihs anu Xurnool.; other (:h -n the

known occur:-'.-m:es oc Lower Cud:: o pah J h.?.vs been brought to

light . i'ur'cfi- t lnvoa'..:.«;• tion i v/i]i reveal thr (Ctu--3.!

cuayc of th- • .<•. inoci-jlie.'-; Lit «i . •: tion to their ;rotc:nLU:-

lite... .....; well a:. otli-T a.3pe«t;j of minorc:li;;..tion.

- 102 -

ACKNOWLEDGEMENT

The authors are highly thankful to Ghri A . C . Saraswat,

Director, Atomic Minerals Division, Department of

Atomic Energy, for all the guidance, encouragement and

support extended for carrying out the investigations.

They are also highly indebted to Sarvashri G. Chakrapani,

H£ndakum?.r, K. 3ubrohmaniam, and Thangoraj of the

Chemistry Laboratory, AKD, Southern Section for the

laboratory support and to Shri ..rjuna P;jnda for his

valuable suggestions.

REFERENCES

King, W. (1872). Kadapah < nd Kurnool formations in the

Madras Preridency, Geol. 3urv. Ind. Men. 8 (1), 320 pp.

Nagaraja Rao, 3.K., Kajurkor, S.T., Ratnalingasi -imy, G.,

ind RavinUra Uabu, 3. (19C7). otruti-jrophy, structure

.inc.' evoluatlon of ttm Cuddapah basin, neol. Soc. India,

Mem. 6, p. 33-06.

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- 104 -

77 ' 70" 79 ' _B0^

MAP SHOWING CUDPAPAH 0ASIN A. P.

50 Scale

91

50 Km

•Guntur

INDEX

11*

16'DD KURMOOLSBO CUODAPAHSIV] NELLORE SCHIST5} GRANITIC GNEISSESg Area cobv Ceochemlcal

BELTESeredsurvey*

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- 106 -

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- 107 -

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HISTOGRAM BASEO ON'U' CONHNI IN W t U WAH R

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- 108 -

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- 109 -

AN ALPHA-GAmA INTEGRATING DEVICE TOR URANIUM EXPLORATION

GIRIDHAR JHA AND P) RAGHAVAYYA *1*1.N. SRINIVASAN AND 5 SHASTRY **

It is often found that location of uranium mineralisationbecomes difficult in areas uhere soil cover is considerable,because of poor gamma ray response. In sycb areas, measurementof integrated concentration of soil gas ^ z R n along uithcumulative gamma dose helps to detect the concealed uraniummineralisation. This combination method uas tried in SinghbhumThrust Belt in eastern India uhere hidden uranium mineralisationuas suspected. Exposure cups equipped uith cellulose nitratefilms used as detectors for measuring soil-gas radon concentra-tion and CaS04 (Dy) thermoluminescent dosimeter for measurementof cumulative gamma dose ware used.

INTRODUCTION

SinghbhuTT) district of Bihar state in eastern part of India

is a treasurehouse of various economic minerals viz. copper,

uranium, iron, phosphate and asbestos etc. In this district,

the Singhbhum Thrust Belt (STB) which extends over 160 Km3 in

length has about half a dozen uranium deposits, uhich include

tuo operating underground mine at Jaduguda and Bhatin. In

1986 - 87 radiometric survey involving measurement of soil-gas

radon and integrated gamma radiation uas undertaken on an

experimental basis for locating a hidden source of uranium in

STB, uhere chances of locating such source of uranium mineral

concentration appeared promising.

SELECTION OF THE AREA

Uhile monitoring water sources, it was found that some

well water and spring uater samples around the village,

Rajdcha in STB, gave radon concentration of 4000 to about

14,000 pCi/l. The soil samples from the same location analysed

16,700 pCi/kg of 2 2 6 R 3 .

* Environmental Survey Laboratory, BARC, Jaduguda Mines*

**Uranium Corporation of India Ltd., Jaduguda Mines.

- 110 -

Sines all these results uera clearly anowloui, it ues decided

to aelect an area measuring about 270,000 m2 in Dungridih-Rajdoh<\

region for soil-gas radon and integrated gamma radiation survey.

Geological setting

Regional geology of Singhbhu* hi a been the subject Batter

of intense study for the past three to four decades. A nuaber of

geologists have studied ths srss and suggested different versions

of geological sequences. The sequence established by DUNN and DEY

still finds acceptance in geological circles. They have divided

ths area into two divisions - north of the thrust belt end south

of ths thrust belt. On the northern slds of the thrust belt, ths

rocks of Chaibasa and Iron Ore stsgss of Iron Ore Series hevs

been reported. On ths southern sitfe of ths thrust belt, rock of

Dhanjorl stags. Iron Ore atega and Singhbhu* granita have baan

dsscribsd. Ths thrust zone is baliavad to have been developed

between Chaibass and Iron Ora stsgs rocks. Tha thrust bslt varying

in width fro* a raw hundred swtrea to'a few thousand us tr as, ax tends

over a length of about 160 kaa, in an arcuate shape (Flg.1).

Tha geological sequence la aa followa.

Slnohbhusi Stratigraphy aftar OUHH

North Slnohbhusj South Sinohbhuai

Chotanagpur granite Slnghbhua granita - dlorita

Oslna lavaa Dhanjori stags - lsvs.qusrtzcongiomarata

Iron ore stage - phylllCes, Iron Ure Stsgu - phylllta,tuff,quartzites, arkorfe,tuff and baalc conglomeratetignaoua rocka quartzita and

baeic IgnaouaChalbaea ataga - ailcs achlat rocka.

hornblanda achlatquart achlat andtuffa.

- 111 -

Tha lithologicel units of the thrust belt are not found in

the areas either to the north or to the south of tha belt. The

priclpal rock types in the thrust belt eru, quartz chlorite schist,

brecciatsd quartzites, basic schists and b^sic igneous rocks. Moat

of the rocks have low dipa. Copper and uranium mineralisations ara

found mostly in the quartz chlorite schist and brecciated quartzites.

Rock formations exposed at some places in area uhera the radon survey

was dona are brecciated quartz!te and quartz chlorite schists. These

rocks strike NU-SE and have dips varying from 30° - 40* in NE direc-

tion.

Soil-gas radon meaaurementa

Conventional radlometric techniques of measuring bata or

gamma ray reaponaaa using CM counters, gamma ray scintillo-maters

and spectrometers ara effective toola of exploration for uranium

mineraliaation, aa long aa a coherent response is obtained from

aurfecs axpoaurea. Target identification ia rendered difficult

uhsn tha aignala ara inadequate. In tha field, bete or gamma ray

reaponee from a eource, ia vary mucn dependent on variable topo-

graphy end tha thickneaa of tha overburden. In these conditione,

any method capable of providing information, about tha extent of

minaraliastlun, depplte depth of overburden la of immenae halp.

Tha method of msuauring integrated concentration of radon

in aoil-gas ualng solid state nuclear track detector (SSNTD) and

gamma dose with Tharmolumlnlacent doalmeter (TLO) ia helpful in

the search of uranium mineralisation, even in areas of incoharant

gamma ray raaponaa (JHA '87). Radon la produced by tha decay of23flradium, a member of the U decay series, which is widely dieti

butad in tha earth1a crust. Radon ia an inert radioactive gat,

- 112 -

which decays with a mean U f a of 5.5 daya emitting alpha parti-

cles. Atoms of radon move long distances froa the site of origin,

both laterally and vertically, through thicK overburden, without

reacting with the medium. The technique of aeaauring integrated

radon concentration and geoaa radiation dose in the soil-gas for

locating uraniua alneralisation relies on recognition of distant

signals in the presence of the background noise.

Waterlsls and methods

Integrated radon concentration, and cumulative gaaaa radia-

tion dose were Measured using an exposure cup. Exploded view of the

exposure cup is shown in Fig.2. Cellulose nitrate fila was the

nuclear track detector used for recording tracks forasd by alpha

particles froa radon and its daughter products. CsSo. (Oy) was the

TLO used for the aeasureaent of gaaaa dosa. The detectors were

aounted on s rectangular aluainiua card placed inalde the cup. A

latex aembrance 100/ua thick was us«d at the other end of the

device to discriminate against the entry of radon isotopes other

than Rn. (3HA 82). for ainiaiaing the effect of aoisture on the

detector, common salt was placed inside the cup as desiccant (3HA*B7),

An area of approximately 1800 a x 150 a was divided into

rectangular grlda measuring 100 a x 50 a, length along NU-SE

direction, which coincides with the strike direction of the rocks

and breadth along* NE-SU direction. Exposure cups were iaplaced

at the interaection polnta of the grids. Pits 15 ca in disaster

wars dug at each sampling point to a depth of 30cm (3HA'87)«

Exposure cups were placed in.the pit with the membrane aide

facing the pit bottom. Pit openings were covered with baked clay

- 113 -

tiles, over uhlch a PVC sheet (thickness 500 /um having radon

permeability co-efficient of 5x10 cm /sec) was spread. Sides

of the PVC sheet uere used for sealing the pit opening uith the

soil obtained from the respective pits. The exposure period

was about 3 to V ueeks. Besides the Oungridih-Rajdohs area, two

more locations - one at 3aduguda nine and the other near Rohinbere,

about 2 km south of Jaduguda end auey from the thrust belt, uere

also surveyed. This uas dona to obtain radon v&lues in a known

uranium deposit and in areas away from the know uranium minera-

lised zone for reference. The Iocstions of the sampling polnta

are shown iivfig.3. At the end of the exposure period, the cups

uere retrived; detectors ware r(moved and cleaned. Gamma doaaa

war* reed using a TLD reader. The SSNTO films were etched in 10%

NaOH solution, at 60*C for two hours, in an Incubator. Cellulose

nitrata layera of the films ware atrlppad from ihe rigid plastic

base and the alpha tracks davalopad in tha film ware countad

uaing either spark counting or mlcroscopa counting techniqus

dapanding on tha track danaity (CD at al 1984). Tha track density

obtained in each M i a waa normalised to 30 days exposura and

than converted to radon axpoaura uaing tha calibration aquations

CE - 20.08 x T0 # 9 8

Uhere C- ia tha radon axpoaura (pCl/l h)

T la tha track danaity (Tracka/oa2)

Tha intagratsd radon axpoaura v»luas wara converted to tha

concentration flguraa (pCi/l) uaing appropriate transforaatlona.

- 114 -

Results and diacuaalona

Integrated radon concentration and gamma dose valuea,

for each aample location are given in table 1. Statiatical

aummary of the data in table 1, ia preaented in table 2. Cumu-

lative frequency plota for the relevant populatione afe alao

ahoun in Fig.4 and 5. Statistical parameters viz. background(b),

atandard deviation( g) and threahold (t) ware calculated from

the equation of the log-probability plot. Threahold uaa calcu-

lated aa the product of geometric mean and a* aquaro of the

gso«etric standard deviation (itPELTIER '69). Fro* tabla,2, it

can be aeen tht't the geometric mean and standard deviation for

background location (RGKIAI8ERA) i« 87.4 pCi/1 and 2.8, respecti-

vely. The threshold for this uorks out to 685 pCi/l.

Considering the concentration of soil-gas radon values

above 685 pCi/l ea anomalous, it is observed that about 195

values frost Dadugude, 15jt and 43.SJC valuaa froai Oungridlh andt

Rsjdoha location* fall in thie category. The log-probability

plot anoua a straight line fit except in Rajdons plot which

ehous breaks. For auch braaka threshold can ba road following

simplified statistical method of UPFLTIER.

Cumulative frequency distribution curva for radon in the

case of Rejdoha shows two braaka. This la an indication of

bluodal distribution, co«prialng of two distinct populations.

By apliting the data at a value taken around the middle of A+6

(800 pCi/l) it is possible to separata ths tuo populations, of

which ths lower ons rsprasants ths background snd ths higher on*

ths anomaly.

- 115 -

From isorad curves presented in Fig.6, it is observed

thst radon peaks appear around sampling points 1 to 3 and 14 to

23 in this area. This observation is also supported by the cumu-

lative gamma dose from the respective locations.

Conclusion

In uranium exploration, radium and radon are well known

path finders, especially in areas where soil covor is considerable.

In the area under study, radon concentration in aoil is about

20 times the background value and exceeda the threshold by a

factor of 3. Radon being a daughter product of radium, its concen-

tration ia controlled by the diatrlbution of radium in the soil.

Concentration of aoil gas radon of the order of 2000 pCl/l obtained

in thia area cannot be supported by the amount of aoil radium

preaant In this region. Hence, there must be • source other than

•oil radium, for such • high concentration of radon to exist and

la indicative of • hidden source of radium, which by inference

point to ur&oium minerallaatlon.

It ia wall aetsblished that water under praaaure can dis-

solve large quantity of radon while couraing through rock forma-

tion a and aoil atrata and theaa water* can trenaport radon to

far off placeaa. Hydrogeologlcal conditiuna in thia area do not

anviaage aucha a poasibillty. Major faults and fractures era

known to give redon anomalies. Examination of outcrop* suggest*

Httl9 poflblolty or a major fault underlying tho r*eton 9nom»ll»»,

The most important point ia that, the** strong radon snooa-

liea are pretent in the SinghDhum thrust belt, where all the major

- 116 -

known deposits of uranium exiat. In fact this Dungridih-Rajdoha

areas is only 1-2 km NE of Naruapahar uranium deposit and about

6-7 km NU of Bhatin uraniua mine. Besides the soil-gas radon

anomalies, water samples collected from springs have also givan

dissolved radon concentration, in the range of 10,000 to 12,000

pCi/l. Soil radius concentration near the spring have recorded

7000 - 16000 pCi/kg. All these signals positively indicate the

presence of urenium mineral concentration in the nearabout region,

probably at depth. Thia areas therefore la most suitsble for sub-

surface exploration by drilling.

Acknowledgement

Authors ere grateful to Hr. fl.K. BATRA, Chairman and Reneging

Director, Urenium Corporation of India Ltd., for hla encouragement

and keen lntaraat In this work. U» are Indebted to Shrl S.D.SOIVLN^

DIRECTOR, Health and Safaty Croup, SAftC /or the invaluable suggea*

tiona and for according permieeion to undertake this work* Thanka

are aleo dua to Br. P,ftvnARKOSE of Environmental Survey Laboratory,

Jaduguda and flr. A.K.SAftAJfCJ of UCIL for their kind essiatance

in the laboratory and field raapactlvely.

- 117 -

Reference*

1. DUNN 3.A. AND A.K. DEY (1942)

•The Geology and Petrology of Eastern Singhbhum and

aurrounding ereea".

memoirs of the Geological Survey of Indis-Vol.69-Psrt III,

2. JHA G et. ml. (1982)

"Radon Permeability of some membr<nces".

Health Physics, Vol.42,No.5,PP 723 - 725.

3. JHA G and n.Rsghsvayya (1963)

"Development of a Passive Radon Doaiaeter".

Proceedings of the Fi f th Netionel Symposium on Radiation

Physics, Nov. 21 - 24, Calcutta.

4. JHA G at a l (1964)

"A Spark Counter for Counting of alpha trucks in SSNTD fi lms".

Bulletin of Radiation Protection, Vol.7,No.1,3an.-March,PP 39-42,

5. 3HA G (1967)

"Development of a passive Radon Oosluster for applications

In radiation protection and uranium exploration".

A thaaia submittad to the Univ. of Bombay for the award of

degree of Ooctor of Philosophy in Physics,

6. UPCLTICR C (1969)

"A Simplified Stat lst lcsl Treatment of Geochemlcel data

by graphical representation",

Econ.Geo. Vol.64, PP 536 - 550.

- 1 IB -

Table - I

Radon concentration and Integrated gamma values in soil-gas ofRajdoha - Dunqrldih

SampleNo. (

1 .

2 .

3.4 .

5.6 .

7 .

8 .

9 .

1 0 .

1 1 .

1 2 .

1 3 .

13b.14 .

15 .

16 .

17 .

1 8 .

1 9 .

2 0 .

2 1 .

2 2 .

Gmmtrm dos«• i l l i re»x30d)

90.4090.70

133.00191.00150.70

134.30102.10105.70112.80116.40107.10192.10181.40

857.00202.10130.00

1057.00908.00

1034.00903.00981.00

-

953.00

222Rn concn.(pCi/1)

986.70756.60

1257.00364.70169.70501.00460.30596.30248.80401.10292.50400.00486.00

1014.10SS5.60806.80404.60

1167.001117.601373.701351.40

-

1891.20

SampleNo.

23 .

2 4 .

2S.

2 6 .

2 7 .

28 .

2 9 .

3 0 .

3 1 .

3 2 .

33.3 4 .

3 5 .

36.3 7 .

3 8 .

3 9 .

4 0 .

4 1 .

418 .

Gamma dose(nilli rexx30d)

1007.00878.00771.00636.00933.00877.00860.00833.00624.00731.00578.00695.00703.00810.00692.00692.00675.00

630.00559.00480.00

222ftn cone.(pCi/1)

1202.80130.60100.40475.80114.8058.20

146.3036.1068.20

530.00350.70362.90455.40

102.30699.00201.00236.00654.80264.00 /

1114.60'

- 119 -

Statistical evaluation of aoll-qaa

222Rn concentration data

Table - I I

Location

figC

Statlaticai information

Porcentile (pCi/l)

95 60 20

ROHINBCRA

3AOUGUUA

OUNCRIOIH

RAJDOHA

87.4

311.1

263.7

602.4

2 .8

2 .4

2 .7

2 . 0

17.6

79.3

56.6

203.90

67.4

249.5

205.5

505.5

208.1

651.5

605.7

1082.3

- 120 -

Fij-1. REGIONAL GEOLOGICAL MAP OF SINGHBHUM

THRUST BELT & ADJOINNG AREAS

TCRTIARr ROCKS.

GRANITES.

OAIMA/OHANJORI LAVA.

ONANJORI OUARTZITB/

CONGLOMERATE.

IROM'ORE STACE ROCKS.

CHA16ASA STAflE ROCKS

THRUST U l T .

- 121 -

. EXPLODED VIEW OF EXPOSURE CUP

0 1 2 3 4c

PERFORATED PROTEgiVE CAP

» . ^ ' »-^ ^ ^ ^ - " ^ - * ^ ^ »•» »^ *-« ^ ^ »

DlSCRtMllMATOR MEMBRANE

^EXPOSURE CHAMBER

TLD AL-CARD

BACK SEC-AA

^hca^zzTZLrnzcaJ

- 122 -

N F,£.3.. MAP SHOWING SAMPLE LOCATIONS

INDE

O SAMPLE STATION

I £

ru

I

a -

2 -

% •

* •

J

3

X •

t

t •

i»i'

t

o

in

* UJ

I

6

a

- 124 -

ff.*T «*.» « .» *• »» N TO «0 tO «0 M tO IS 1 1 I OS Ol ! | MS

• OUNCRCM

• RAJOOHA

CUMULATIVE PERCENT MORE THAN STATED

LOG PROBABUTY PLOT OF SOIL-GAS 222R«. CONC.

OBTAINED FROM JADUGLOA.DUNGROH - RAJPQHA.

- 125 -

PLAN SHOWING CONTOURS OF RADON CONCENTRATION IN SOIL-GAS

0 10 100 IW 100m

LOCATION

CQNTOUR WTEFWL 100

- 126 -

GEOSTATISTICAL STUDY OF BHATIN ORE DEPOSIT - A CASE STUDY

C.V.L.Vajpai and P.P.Sharma Uranium corporation of IndiaLtd.

Bhatin is a small uranium deposit being worked by UCILin Singhbhua district of Bihar* Large dispersion of R.O.N.grades have been causing a concern to a great deal fromquite some time. Conventional technique adopted for rwrvsevaluation lacked accurate prediction of grade fluctuations.Geostatistical technique is used for reserves evaluation.Error involved in estimation is calculated. Attempt has beenmade to study these wide variations in predicted and achievedgrades. Estimated grades of the deposit by both the techniquesare compared, best estimator for grade of unknown mining blockis evaluated. i

Geostatlstlcs in uranium ore reserve estimation

Once an uranium deposit has been located in a certain area,a regular grid pattern of boreholes will be made and the gradeis determined by logging of the boreholes and radlometricmeasurements and chemical assay of core samples of each bore-hole and the volume of the ore body,grade,accumulation (gradetimes the thickness) and other essential parameters are roughlydelineated. Using these values, so far, it was customay to useclassical statistical techniques(for example polygonal orinverse distance method and Slchel's *t* estimator) to obtainore reserve computations. However these methods mr* not precise.

- 127 -

The Inadequacies are well Known and fundamental objections

are that the procedures for assigning values to the chunks

or ore body are quite arbitrary and without a sound theo-

retical basis. The so called 'principle of gradual change'

and the 'rule of nearest points* are not exactly based on

any mathematical principle. The methods can be biased and

the estimated procedures do not usually include a method of

determining the precision of the estimate. In recent years

geostatlstical methods and Kriging(as developed by G.flatheron

of Fontainbleau School) are increasingly used by mining

engineers and practising geologists for interpretation of

spatial data and for arriving at optimum estimates of ore

reserves, especially for deposits of gold and uranium. These

methods do not have the deficiencies mentioned earlier and

are based on firm theoretical concepts. Without going into

mathematical rigour, the philosophy of the method can be

briefly outlined as follows :

Concepts of Geostatiatlcs

Ceostatlstics accepts the concept that each point in tbe

deposit represents a sample from some distribution function,

but the distribution at any point may tiffer completely

from that at all other points In the deposit in its fora,

mean and variance. If the difference in grades is taken

between two points (P^ and p1 say) separated by a distance

h, than this difference will ba a variable that follows

a distribution dictated by the distributions at each of the

two points. If we tax* another pair of points the same

distance apart and having the same orientation, the difference

in grade between these points will also have a distribution*

Geostatlstics assumes that the distribution of the difference

in grade between two point samples is the same over the

entire deposit and that it depends only on the distance

between and the orientation of the points.

- 128 -

IfAg(i'} are grades at ?- and P- separated by a distance

IQI alcn£ x direction, where i • 1 • h and 1 assumes

from 1 to n* Tbe successive differences square

can be averaged as C^ (^ ' ft

This is danotcd by vari.o&ram function 2./1[hJ (where the

Tactor 2 is a natter oz .aatneajatlcal con»eniencejGrapining

ot-this function is done in tne usual saaner, with values

of the function plotted on the Y-axis ana the distance h on

the x- axis.

M OF THE HIBZRAL JEPOSIT

Ones the volose and ohapa of an uranius ore txxy :u* been

defined and the aaount cf available data is «ii9iignflt is

possible to carry out a detailed and sufficiently precise

grade estimation of tne various blocks into wnicii *h» ore

body oas been suitably divided acoording to the stoping

dMlgn* 3b» prooadure for saklng a'teostatistical or* rsssrv*

ostlaation can be divided into two parts. Tea first is the

investigation aad aodellng ot tbm physical and statistical

structure of tha orm body* Concepts of ooatinulty aad struc-

ture la the deposit are —bodied in varlofrucs that art

constructed during tbe first step* The second stage of the

procedure la tbe estlaetloai process Itself- Krlginc-whlon

depeada entirely on tbe varlograae draan during tbe first

stage* the riguree exparl—iltaiiy obtained, provide tbe

polnta of tbe experimental varlogrssi V*(W • ay repsatlafthe saae procedure in other dlreetloaavsey csst-«estf aortb*south north* east to south-west and/or north* weet to sooth*east, we get different varlograas. Vhlle these experimentalvariocraaa may help In deteralaint the atruature of •

- 129 -

deposit and the behaviour of grade variations it must

be related to some theoretical model if conclusions

are to be drawn or to make estimations for unsampled

areas in the deposit* The commonly used models for

theoretical semivariograas are :

(1) the spherical or transitive model also called

Matheron scheme.

General shape and its equation isl shape q n is

V ( N = C o -r C fib. — hi _ ~~\ W*«-:-. h 3 - -

- G

a is called the range, c is the nugget effect and

Co • C is the sill.

Points farther apart than distance 'a' are unrelated

or Independent of on* another*

Krlgln* t The procedure, which yields best linear

unbiased estimation variance for the grade of a given

block and data configuration is known as Jtriging.

Kriging uses the property of the variogram(which

describe the spatial variability), and selects the

weighting factors which minimise the estimation variance.

The estimation variance can be obtained from the follo-

wing Kriging system equations i

- 130 -

grade or the block, X\ weighting coefficientsof the sample grade gi

^ ~ (condition for unbiased estimation) -

^'O'?)/ *i ft* v) are *ne average values obtainedfroa the variogram of the block configuration, jX-

Lagrange factor estimation variance

" i v) - r~{ v v ; + > < - . . - 7

For sinple geometrical arrangements charts are avail-able for the calculation of average variogram values

Y •!?,}). 1\\v)

Y t V, vy for various models when this is notpossible, the equations are solved to get the w^ghtingcoefficients by suitable computer programmes and thenthe block grade and associated estimation variance arecalculated. In the case of uranium where radiometriclogs are used, cokriging can be made between chemicaland radiometric grades.

Case Study ;

fihatin uranium deposit is consideredto be a structurally controlled hydrothermal deposit. Theprincipal structural controls are the shear planes whichstrike approximately NV - SE and dip to the NE. There aretwo lodes (1) the main or the hang wall lode which extendsfrom east to west for a distance of 400 metres and(2) the foot wall lode which is much shorter(approximately150 metres long ).

The thickness of -r.2 ore bodies as also their grade

are variable while the foot wall lode is 1.2 to 2.5

aetres wide throughout, the hangwall lode varies in

width from 2 to j metres at the extremities to over

b metres in the centre. Further, there are evidences

of post mineralisation structural disturbance in flhatin

like cross - folding and strike slip faults, as seen

in the mine openings.

(2) The exploratory and development drilling were

carried out by AMD both along and perpendicular to the

strike of the ore body, with 80- 180 metres and 80-210

•etres apart respectively(Fig 5). The individual samples

from these holes were in lengths oi 15 cms.

(3) The statistics ox the 417 individual core samples

have yielded an excellent lognormal distribution

(fig 1, 2, 3 & 4) and have indicated the presence of a

unique population of accumulation values with a logari-

thmic variance of 0.52 and * logarithmic standard

deviation of 0.72.

(4) The geo«tatistics of these individual samples have

revealed that this ore body is of the transitive type as

shown by fig 6 ( Range of variograu • 52*5 cms,

Nugget effect - 5 x 10"* and sill - 20 x 10~4 )

(5) Variogram of accumulation valves along the strike of

the ore body Is random type (fig 7 )• The result of this

type is rarely seen in ore bodxes untill and unless the

mineralisation is highly heterogeneous one. Arithmetic

average grade of the deposit is estimated at 0.053 %•

- 132 -

(6) Slchel «t« estimator has yielded 0.051 % as the best

estimator for the grade of the deposit. At 95 % confidence

lower end upper rounds of «t» are 0.039 % and 0.067 % and

Individual assay values as 0.017 % and 0.150 % respectively.

(7) The statistics of the 84 individual underground

channel data have yielded lognormul distribution (fig 8 & 9)

with a logarithmic variance oi 0.12 and a logarithmic

standard deviation of 0.35 •

(8) The variogram of underground channel values is

transitive. It has revealed continuity for a distance of

10.5 aetres (fig 10).

(9) Kriged grade of the mining blocK of ore is given

by (fig 11).

Otl'j.t 0. 0 Z {

To summarise the s ta t i s t i ca l studies have brought out :

(1) The distribution of uranium Is lognoraal in Boatin •(2) Lower 95 % confidence level includes over 20 % values

below cut-off grade.(3) Variogram for accumulation for the uranium minerali-

sation in Bhatln i s of the transitive type and thecontinuity i s maximum along strike and least across i t .Strike variogram of bore bolt data depicts insufficientgrid s ize used for evaluation of the nawvea of the*deposit. Waste zone i s found to be increasing at thetime of mine development.

- 133 -

(5) Kriged estimate is found to be the best estimatefor underground mining blocks. It has helped to controllarge R.O.M.grade fluctuations.

Acknowledgements ;

Authors are indebted to Shrl S.Shastry,Additional SuperiTtandent(Otology),UCTL, for his valuablesuggestions and encouragement in carrying out this worK.

Thanks are also due to Shri J.L.Btiasin,Chairman and Managing Director, UCIL, for his supportand interest for this type of work and bis kind permi-ssion to present this paper at this symposium.

- 134 -

R e f e r e n c e s

BLAISE, R.A. and CARLIER, P.A.,(1968) Application ofGeo-statistics in Ore Evaluation, Canadian Inst.offlin. & net, Spl.vol.9,pp.41 - 68 .

BROOKER, P.I., (1979) Krlging £ & MJ,Vol.180,Mo,9,pp.i48.

CLARK, I.,(1979) A review of the theoretical foundationsof geostatistics and the practical methods of constructinga ssnivariogram, E & MJ, Vol.180,Mo,7,pp.90.

CLARK, I.,(1979) How to fit a sioplistic model to anexperimental semivariogram, E & MJ, Vol.180, Ho.8,pp.92.

DAVID, M. , (1977) GEOSTATITICAL ORE RESERVES ESTIMATION,Elaevler Scientific Company, Amsterdam.

DIXQM, W.J. and MASSEY, F.S.,(1957) INTRODUCTION TOSTATISTICAL ANALYSIS, Hograv Hill BOOK CO., N«W Yor*.

JOUftifcL, A.G.,(i979)G«ost8tistica& Siaaulatlon methods forExploration and nine Planning E A MJ, Vol.180,No.12,pp.86.

PARKER, H., (1979)The volume variance relationship : A usualTool for nine Planning, E A rtJ, Vol.180, No. 10, pp 108.

RENDU, J.n.,(1980) A ease study ot kriging for Ore valuationand nine Planning, E & H J vol.181. No. 1,pp. 114.

ROYLE. A.G.,(1979) Why Gaostatistics ?, E & rtJ, Vol.180,No.5 t PP.92.

SANDEFUR, R.L. and GRANT, D.C. ,(1930) Applying Gsostatisticsto Roll Front Uranium in Wyoming, E & MJ, Vol.181, No.2,pp.90.

- 135 -

SRI NIVASAN, M.N. and VAJPAI, C.V.L.(1986) Exploration

and evaluation fcr uranium minerals in Bihar presented

in symposium on • Mineral Exploration,Challenges and

Constraints • held at Patna, Blhsr Dec11986.

VENXATARAKAK, K.,SHARMA R.N. and VAJPAI,C.V.L.(1971)n An attempt in the application of the Geostatistics to

the uranium mineralisation at Jaduguda ", Symposium on

uranium held at Jaduguda , Bihar Jan 1971.

V2NKATARAHAH, K., VAJPAI, C.V.L.(1975) A statistical

approach to the study of uranium mineralisation at

Jaduguda , District Sir:ghcbua(Bihar; Jouruel of tne

Geological Society of India Vol.16, No.;}, 1975 PP 354-360.

cvlv

.

I :

SO

XGO

r,< < !

-i

3! I

*lL_Jc•

r s

*(1

A

<CO

- " • . \ . — . • ; • * ~

••-, .:.-•?

r_> ' .•rtt.- :•••» ' . • . « . •

137 -

...i"

100

SO

LOG-ASSAY-FREQUENCY HSTOGRAM- OF B-H-DATA

NUMBER OF SAMPLES *» 417 .. ,

140

K>

0

A\

V

W . M- 12 M M 15 Ifi l> 18 13 20 . ?l r

LOG Af.SAV ->-

1 • .

i .

. I . . . . .

' ' ' ''i' '

• • • ; : : : . ' • . I

.'. i '• -:'.';... : ' '

2-3 24

- 138 -

——-f

- -K-O-

•iSO

ISO

•tic

Q T •_= - vf - r , . - i - - - g & a r - = - . : . . , H _ - , j ; ^ , 4 . i . .

T O T A L N o OF JS^MPLES ' 4 1 7 -

• . • • • • • • • : y - • ; • '

- • : • / - — - —

• •• 9

I• - • - —HO~i

(

•:*?!

iI3C-

-• *?

•no

uo

t9

<< -C70

• M O

osa

.-- -CiO

-020

•01?

Fiq.3

r!

_ j . c .

t.0 • *6

-O--H

- 139 -

: : _ /

'•3

i-c.

PLOTJOF ^UHULAra>"E PPEQUEKCY PERCENTAGE Vs LOG ASSAY

.— J _ ^ _ : _ CUMULATIVE/-* i.

" T O T A L No-OP" SAMPLES = 417 •"•""• :

- uo -

\

c °

OC

oo

5 ' -^CM'

• ORE HOL.E LOCATIOM• HAT IN 'MINI

SCalfti- V. 4OOO

Fi

- 142 -

• _ 1 . •• I

SEMI \»RK)GRAM OP ACCUMULATION VALUES /

OF ORE BODY C+d5m LEVEL)

. No Of BOREHOLES =3

No OF SAMPLES* 3

THE STRIKE ..

XCJUO

2MXV0

26OHG'

240X19*

it 143* |C5

JO

40II10*

20X10

KC IcC W

DISTANCE IN METRES->

if.

F/j.7

- U3 -

ACCUMULATION FREQUENCY ' HISTOGRAM OF CHANNEL DATA

NUMBER OF SAMPLES = 84

tc

12

2

2

O0 2

1 1

i i

4C 12 Id 16 16 ' 20 22 24

ACCUMULATION

20

M

v- M

— Ul

IIit--

..:«6

. • *

- 144 -

LOG - ACCUMULATION FREQUENCY HISTOGRAM

NUMBB'liQF, SAMPLES_-

OR -.CHjANNEL

i i ; •'

I 1»S0

JZ1__.3-80

LOG -ACCUMULATION

400 405 4-15 .

.1 . .-.:.•.

:.1:;ll

485. 4-30 i

- 145 -

w.-r"

t ; ,:.-..• j

S I " ? 1 \ I ' M j ; ' • >• • •

. o

40

I!! ME.TRL-S • • -

f ' - r' * . l ?

1

u>o n o ieo wo 200

ui

CO

</)UJ

a

I

0)

I

a:

a9Q:

S E S S I O N II B

ANALYTICAL TECHNIQUES IN URANIUM TECHNOLOGY - I

Chairman : Dr. tf.C. JAIN, BARCReporteur: Shri V.N. KRISHNAN, BARC.

- 147 -

URANIUM ANALYSIS USING AN ON-LINE BACKGROUND CORRECTION PROGRAMWITH CARRIER DISTILLATION TECHNIQUE BY A COMPUTER CONTROLLEDEMISSION SPECTROMETER

R.K.Dhurawad, A.B.Patwardhan. V.T.Kulkarni, K.RadhakrishnanFuel Reprocessing Division, B.A.R.C., Trombay

Bombay - 400 085.

SUMMARY

The paper describes an on-line background correction anduraniua monitoring (due to occasional matrix excitation)in acomputer controlled Direct Reading Spectrometer during theestimation of a large number of impurities in uranium productused in nuclear facilities.

The influence/interference of the background on theanalytical lines becomes important when low detection limits areto be achieved. Commercially available softwares for automaticbackground correction (ABC) are suitable for InductivelyCoupled Plasma or Spark sources where one can have a continuous,flow of the sample introduction (without much restriction on timeof exposure). However, in the case of carrier distillationtechnique automatic background correction cannot be applied dueto limitations of exposure per charge. In the method discussedhere, a background channel located at an appropriate position inthe spectral range is monitored simultaneously along with otheranalyte channels. The background at analyte channel is computedfrom the intensity of the background channel and is automaticallysubtracted from the intensity of the analyte signal. In addition,a uranium channel which is used for Monitoring a weak line ofuraniua (286.567 nm ) is incorporated to measure the amount ofuranium getting into the arc. When intensity of uranium lineexceeds the predetermined value, the data will be rejected by theoperator. This method is in routine use over a few years for theestimation of 2t impurities in uranium.

INTRODUCTION

For quality control of uranium required in nuclearfacilities, carrier distillation technique was developed byScribner and Mullin in 1946 (1). This method involves 1)conversion of sample matrix to a form having low volatility 2)addition of a small amount of selected volatile "carrier" and 3)partial distillation of the mixture in a d.c arc under optimisedconditions. The limits of detection for a majority of elementsare in the range of a few parts per million. For B and Cd it isnecessary to ensure that the detection limits are 0.1 ppm orbetter. The influence/interference of background on analyticallines becomes especially important at these levels.

Until recently the impurity analysis was carried out in ourlaboratory using photographic imaging and densitometry. It was atime consuming process, with the advent of new technology i.ephoto-multipliers, microprocessors and computers, the

photographic detection has been replaced by Direct ReadingSpectrometers. Pepper & Blank (2) have reported use of DirectReader for carrier distillation using exposure control unit fordifferent elements.

Background correction in Direct Reading Spectrometers forspark and ICP sources have been reported in literature (3,4)using two methods namely refractor plate technique and movingslit technique. Both these techniques essentially require asteady and continuous source over a total period of exposure.This requirement can be met in the case of ICP as well as sparksources. However, in the case of carrier distillation techniquethe sample is not continuously injected at constant rate butinstead a limited quantity of sample blended with carrier loadedon the electrode is consumed during a short period of a fractionof a minute. Moreover different elements have differentvolatilization characteristics depending upon the nature of thesample, atmosphere and electrode material etc. Thus thelimitations of the sample amount and exposure time in the case ofcarrier distillation method are some of the major factors to betaken into consideration. It is not possible to apply the samemethod of background correction as employed in ICP or spark.

In some laboratories background correction in carrierdistillation technique is carried out by subtracting knownsignal which is arrived at by measuring the signal of pure matrixat all the channels and then computing the background correctedintensity (BCD for each channel (element). This is somewhat anarbitrary subtraction.

BCI « I - I t

element blankIn the present method, the approach is unique and is totally

different which can be termed as on-line background correctionwith Direct Readiang Spectrometer. In this approach, a backgroundchannel is located at an appropriate position in the spectralrange. Intensity is read at this channel alongwith othere alytical channels. This intensity is used to compute thebackground of different analytical lines by multiplying intensityof background channel with pre-determined factors.

BCI • I - I • RElem.channel Dkg. channel

Where *R* is pre-determined factor for different analyticallines which is calculated as explained below.

In a given line, if IBL and IBR are the backgroundintensities to the left and to the right of the line respectivelythen the average background is represented by

Average Background - (IBL + IBR) / 2

- 149 -

Por each analytical line K is calculated using the formula

( IBL + IBR ) / 2K • ————————————————————~

iBackground channel

IBL and IBR are found by reading intensity of purematrix by moving the entrance slit of the spectrometer.

RESULTS & DISCUSSION

Table I gives the observed and computed backgroundintensities of a few selected channels .Thus incorporating allthe constants evaluated for different elements in the equation,uranium samples were analysed. Table II gives the analysisresults of two synthetic standard samples. The results are ingood agreement.

The degree to which uranium interference is avoided is alsoan important factor in carrier distillation technique. To keepstrict control on exposure, we have adapted a system wherein achannel for a weak uranium line (286.567 nm ) is monitoredalongwith analyta lines and its intensity is measured. The cut-off intensity is pre-determined which is 1000 counts. Anyexposure exceeding this intensity is rejected by the operator.The results of the exposures exceeding this intensity limit havebeen found to be of the order of 3 and above.

All these results show the usefulness of this new approach.This method is in routine use for a few years for the analysis ofuranium samples.

ACKNOWLEDGEMENT!

The authors wish to thank Shri A.M. Prasad, Director,Reprocessing Group and Shri H.K. Rao, Head, Fuel ReprocessingDivision for their keen interest and encouragement during thecourse of the work.

REFERENCES

1. B.r. Scribner and H.R. Mullin, J. Research NBS 21, RP 1753(1946)

2. C.I. Popper et al, MLCO-1071 cat. UC-4 (1970)

3. R.W. Spillman and H.V. Malmstadt. Analytical Chemistry, <U,P.303-311 (1976)

4. V.A. Passal, SCIENCE, 222 p.183-191 (1978)

- 150 -

TABLE I

Observed and computed background intensities

observed computed observed computed

Element B Element Cd

455

422

434

452

447

517

553

662

696

559

419

414

389

478

437

437

497

513

527

531

293

257

273

311

3tl

347

407

399

387

365

267

248

255

265

257

250

31*

308

298

29t

Element : Co El •suit : Zn

1199

1381

1578

1529

1393

1159

nee1280

1302

1122

1443

134*

1377

1434

1387

11M

1075

Ilt6

1151

1113

386

359

369

369

404

472

385

510

417

417

377

350

359

374

362

352

362

347

405

412

- 151 -

TABLE II

(All values in ppm on uranium basis)

Sample 1 Sample 2*

Element Amountadded

Amountdetected

Amountadded

Amountdetected

Al

B

CO

Ca

Cd

Cv

Cu

Pa

MO

Mn

Ni

Pb

Zn

It

5

2

10

20

10

10

50

25

1*

5

It

20

6.8

4.9

16.7

14.7

25

9.1

9.6

50.8

2%

9.2

4.2

9.5

22

0.1

1

-

0.1

10

2

-

-

1

10

2

—

0.13

2

-

0.16

8.10

2.0

-

-

1.3

9.9

1.5

* Sample 2 is a R«f. standard wherein some of tha impuritias likeAl,Ca ate. ara not addad.

- 152 -

DETERMINATION OP TARACE METALS IN URANIUM OXIDE BY 1CP-MS

S.Vijayalakshmi. R.Krishna prabhu. T.R.Mahalingam and

C.K.Mathews.. Radiochen'stry ProgmmmB,Indira Gandhi Centre

for Atonic Research, Kalpakkam 603102, Tamil Nadu (INDIA).

Inductively coupled plasma mass spectrometry

(ICP-MS) is fast emerging as a sensitive multielement

technique with detection limits below ng/ml levels.

Excellent reviews have appeared in the literature in the

recent past.(1)r(2> This paper describes the method that

we developed and standardised in our laboratory for the

determination of a number of impurities in uranium oxide.

Conventionally the analysis of uranium oxide is carried out

using optical emission spectrometry. Apart from

intrinsically low sensitivity, the technique suffers from

severe spectral interference caused by the complex spectrum

of uranium. Hence either the carrier distillation

technique"' or matrix separation using solvent

extraction*4* or ion exchange*5* are adopted. In contrast

the uranium spectrum obtained in ICP-NS is quite simple.

Apart from the two singly charged isotope peaks at masses

235 and 23$. oxide peaks at 251 and 254. and doubly charged

peaks at 117.S and 119 are the only peaks associated with

uranium. The oxide peaks do not interfere with any of the

impurity isotopes. The doubly charged peaks are isobaric

only with two sinor isotope* of tin which poses no problems

as alternate more abundant isotopes of tin are availbale for

analysis. In the Method developed in our laboratory, uranium

oxide was dissolved in nitric acid and the uranyl nitrate

- 153 -

solution was directly aspirated into the ICP. The

precision, accuracy and the detection limits obtainrd are

discussed in this paper. For achievinf very low detection

limits in the ppb levels. matrix separation was required.

For this purpose a solvent extraction procedure was

employed.(6>

Instrument used: Elan ICP-MS model 250 ( Sciex.Canada) was

used. The instrumental conditions are listed in table 1.

Experimental:

The effect of uranium on the various

analytes' signals was studied upto O.lfc (w/v) of uranium

table 2. Uranium was found to have a suppresion effect. A

concentration of 0.05 X of uranium was chosen as the

optimum concentration level to work with. To take car* of

the instrumental drift and to Improve the precision of the

measurements. Ga.Sb and Tl were used as internal standards

for the low. medium and high mass elements respectively. It

was made sure that the isotopes chosen for the

measurements were free from isobaric interference.

Multiple standard addition technique was adopted to take

care of the matrix effect. To check the accuracy of

the method an IAEA sample of uranium oxide (SR-C4) was

analysed.

Solvent Extraction:

1 gram sample of uranium oxide was dissolved in 10

ml of nitric acid (6H). and the uranium matrix was

selectively extracted by solvent extraction with 60 X TBP in

- 154 -

^. The aqueous phase was found to have only 10 to IS ppm

of uranium which was not found to have any effect on the

analyte signals. Hence the aqueous phase was directly

analysed by ICP-MS us ins calibration taken with pure

multielement standards.

Results:

The detection limits ( calculated on the basis of

three times the standard deviation got fro* twenty four

measurements of the blank) of the direct method for the

various analytes were found to be in ppa-sub ppm levels

(Table 3), which is adequate for most of the common

impurities. Our results of analysis on IAEA reference sample

compare reasonably well with the certified values fiven by

IAEA. The results of the analysis of a uranium oxide sample

after solvent extraction of uranium are given in table 4. It

could be seen that the detection plaits are in the 0.S to 10

ppb levels and the precision around 10 * rsd.

Table 1

INSTRUMENTAL PARAMETERS

Nebuliser pressure

Nebuliser argon flow

Coolant argon flow

Auxiliary argon flow

Plasma power

Reflected power

3t psi Sampling depth s 23 mm

0.4 lpsi Measurement time : 0.25 sec

12 1pm Measurement/peak : 3

2.4 lpm Repeat integration : 8

1.2 Kw

5 Watts

- 155 -

Table 2.

Matrix effect of uranium.Cone: Percentage suppression of the analytical signalof uranium Li Cu In Ce Ho Bi

1000 ppm500 ppm200 ppm100 ppm

866740*

865929*

75456*

704612

-24

6538

-30

6625-24-50

* - No suppression.

Table 3. Direct nethod

Element

AgCdCrCoCuInMgMnMoNiSrTi

Elemi

CeErDyLaNd

Table

ElNM

BaCdCrCoCuMnMoMgNiPbVZn

int

4.

Wit

Con.inIAEA

sample

<0. 9<1.15.94.38.6

<0. 5<4. 914.311.413.8<0. 4<1.9

Det

Solvent

Con:i nIAEA

•amplelPP»)

0.20.325.54.95.814.111.63.713.82.20.74.2

RSD

•

00000

Element Det

LaCePrNdSm

_—12613

—735

—

y. IAEA DelCertifiedrange

3.4.4.

14.9.8.

limits

.3

.6

.7

.5

.6

103

054

extraction

. l i

0.0.o.0.0.

RSD X

61253312115.55667

imi tc

0010020005005003

(ppm)

_—- 5.0- 4.3- 6.7——- 18.0- 16.8-14.1

—

El em

GdPrSmHoYb

method

IAEAovera1)median<PP*O

—3.64.25.015.312.8

11.40.95

—

:. limit(ppm)

0.1.2.0.3.0.4.0.1.3.0.1.

»nt

914565960149

FBR Speci-ficationslimits (ppm)

201300200100

150200200500150100

Det.1imiti

1.80.10.20.41.5

IAEACertified 1

3.4.4.14.

0

Element

GdDyHoErYb

9.

8.

range<PP»)

10305

4.32

—- 5- 4.3- 6.7- 18.0- 16.8—- 14.1- 2.15

—

Det.limit

0.0020.0020.00020.00090.002

i

Det.imite<PP«)

0.0.0.0.0.0.0.0.0.0.0.0.

00700200800040190009003028002004013032

- 156 -

References:

1. D.J.Douglas and R.S.Houk.. Inductively coupledplasma^ mass spectrometry. Prog.Analyt.atom.spectros.Vol-8. pp 1-18. 1985.

2. G.M.Hieftjc, and G.H.Vickers. Developements in plasmasource/mass spectrometry. Analytica Chimica Acta. 216.pp 1-2-4. 1989.

3. A.G.Pa«e etal.. Estimation of metallic impurities inuranium by carrier distillation method. BARC-862. 1976.

4. A.G.I.Dalvl et al. Chemical separation andspectrofraphic estimation of rare earths in (UPu)02:Talanta. Vol.24, pp 43-45. 1977. z

5. B.D.Joshi et al.. Anion exchange separation andsttoctrofraph1c determination of rmrm earths in

flutonioum with LiP/AgCl carrier. Anal.Chim. Acta,7. 379-86. 1971.

t

6. T.R.Banfia et al. Spectrochemical determination oftrace metals in uranium. B.A.R.C - 950. 197S

DEVELOPMENT OFFOR

- 157 -

FLOW INJECTION ANALYSIS TECHNIQUEURANIUM ESTIMATION

A.H. PARANJAPE; S.S. PANDIT; S.S. SHINDE; A. RAMANUJAM;R.K.DHUMWAD

Fuel Reprocessing DivisionBARC,Bombay 400 085

Flow injection analysis is increasingly used as a processcontrol analytical technique in" many industries. This paperdescribes the development of such a system for the analysis ofuranium (VI) and (IV) and its gross gamma activity. It isamenable for on-line or off-line monitoring of uranium and it3activity in process streams. The sample injection port issuitable for automated injection of radioactive samples. Thepeformance of the system has been tested for the colorimetricresponse of U(VI) samples at 410nm in the range 40 to 350mg/mland U(IV) samples at 650nm in the range 15-120mg/ml in nitricacid medium using Metrohm 662 Photometer and a recorder asdetector assembly. This technique with certain modifications isused for the analysis of U(VI) in the range 0.5-4mg/aliq. byalcoholic thiocynate procedure. In all these cases the precisionobtained was found to be better than +/-1.5X. With Nal well-type detector in the flow line, the gross gamma counting of thesolution under flow is found to be within a precision of +/- 5%

I. INTRODUCTION

Flow injection analysis(FIA) is a simple and eleganttechnique which finds increasing applications as a processcontrol analytical technique in many industries. Ruzicka andHansen were the first to use the term Flow injection analysis(1)). In general it involves injection of a sample aliquot intoa steady flowing stream of reagent and passing this reagent-sample mixture through a suitable detector. FIA is a flexible andconvenient technique which can be adapted for continuous processstream monitoring with good precision. In a laboratoryenvironment, such a system can handle many samples vmry quicklyand because of the flexibility, the same system can be modifiedto carry out various analyses.

This paper describes the development of such asystem/technique for the analysis of uranium (U(VI) and U(IV))and its gross gamma activity. It is amenable for on-line or off-line monitoring of uranium and its activity in process streams ofuranium extraction and purification plants and in fuelreprocesing plants based on Purex process. The sample injectionports are suitable for automated injection of radio aotivesamples

II. REAGENTS AND CHEMICALS

Uranyl nitrate solution in 0.1M nitrio acid andelectrolytically generated uranous nitrate in 1.0M nitric acid

- 158 -

and 0.1M hydraaine were used and their concetrations wereestimated using standard analytical procedures. Alcoholicthiocynate reagent containing stannous chloride and ethylacetatewas prepared as described elsewhere(2).

III. DEVELOPMENT OF FLOW INJECTION ANALYSIS SYSTEM

The technique for the estimation of U(VI) in the range 40 to350 g/1 involved injection of the sample aliquot at a steadyrate into a steady and continuously flowing stream of dilutenitric acid and measurement of its colorimetric response at 410nmwith a suitable photometer. For monitoring the gross gammaactivity, the same diluted solution was passed through a gammacounter. The modular configurations used are given in Fig.l. Withminor modifications in the sample delivery unit and injectionport assembly, this technique could be used for on-line or off-line analysis of uranium. The method used for the analysis ofuranium in the range 0.5mg to 4mg was based on colorimetry at420nm, with alcoholic thiocynate as the chromogenic reagent. Tocarry out this analysis, the FIA technique was modified so that a

closed loop" or" stopped flow' procedure (3) could be used.Here, the sample and the reagent are either mixed thoroughly atthe injection port or circulated in a closed loop till a steadystate response is achieved at the detector. Though any of theabove two configurations could be used for the colour measurementof U(IV) at 650nm after dilution with nitric acid, the tirstconfiguration was tested.

The FIA system can be divided into three modules: a)delivery units for maintaining steady flow of the reagent ( andsample, if required) b) injection module for introducing sampleand c) detector module. The various units employed in the abovementioned configurations are detailed below.

(a) Reagent Delivery Units

For direct colorimetric estimation of uranium, amicroprocessor controlled Multi Dosimat Unit (Metrohm, Swiss,model No. 665) was used to deliver 0.1M nitric acid at aconstant rate that can be adjusted to any desired value. Thisunit was used with 10ml volume burette (No.6.3007.210). Theaccuracy of the volume delivered is +/-0.2X. For delivering thealcoholic thiocynate roagent, a piston pipette (Repipet.USA )with a capacity of 10ml/stroke was used for delivering 10 mlreagent at a time with a volume accuracy of +/-1.5*.

(b) Sample Delivery Units

The sample aliquots were delivered by another Multi Dosimatwith one ml capacity burette (no.6.3006.113).With this unit, anyaliquot within one ml could be delivered at a constant rate thatcan be adjusted. The accuracy of the volume delivery is +/-0.3X.

For on-line monitoring purpose with a dedicated FIA systemincorporating the Multi Dosimat Unit, the sample solution was

- 150 -

sucked straight into the one ml burette and delivered Into theinjection port assembly via a permanent microbore tube connectionand this operation could be repeated any number of tiroes.

For handling radioactive samples during off-line analysis ofuranium, the delivery mode of the Dosimat was modified to avoidcross contamination. In this case, the burette was connecteddirectly ( without, going through the three way valve) to adisposable polypropylene pipette tip of one ml capacity through aflexible narrow bore teflon tubing of appropriate length ( 1 mlcapacity). The burette and the teflon tubing were filled withwater such that during sample sucking and delivery, the watermoves back and forth within the teflon tubing thus reducing thevolume of the air pocket above the sample in the tip. The sample(lml) sucked into the tip was delivered into the desiredinjection chamber either as a continuous stream or as distinctaliquots of smaller sizes at the chosen flow rate. In the latercase, usually the first aliquot was rejected. After pipetting,the tip was disposed off. The delivery volumes have an accuracybetter than +/- 1.0%. The derails and reliability of a similarunit are described elsewhere (4).

(c) Injection Port Assembly

Injection port assemblies form a crucial part of the FIAsystem and in the present work, they had to be compatible withsafe radioactive practices. Two different types of injectionchambers as shown in Fig.l, were chosen for testing. Of these,the first one was a fully closed system meant mainly for on-lineanalytical applications and the second one was an op«m-cup typesuited for automated off-line analysis in laboratories.

(i) Fully Closed Injection Chamber

The fully closed injection chamber assembly was made up ofa pyrogen device used for glucose drip adjustment in hospitals,with additional penetrations for sample entry via microboretubing and a vent line for adjusting the pressure build up andfor controlling the liquid level inside the injection chamber.Shortly after starting the reagent flow, the sample was injectedinto the chamber at a steady rate where it got diluted andtransported via narrow bore tubing to the photometer. The reagentand sample delivery points in the injection chamber were locatedin such a manner that during their delivery a good miking of bothtakes place in the chamber Itself. Being a closed system, thesample-reagent mixture passed through the narrow bore tubingunder slight pressure exerted by the Multidosimat units.Sufficient tubing length (90cm long/1 mm bore) Was provided toget adequate mixing before the mixture reached the photometer.

For on-line monitoring, the sample delivery tube of theDosimat was directly connected to the injection dhamber. Wheneverthis injection port assembly was used for off-line analysis, the

- 160 -

sample was injected through the septum in the port using thepolypropylene tip.

(ii) Open-Cup Injection Chamber

Inorder to avoid the necessity of injecting the samplethrough the septum during off-line analysis, an open cup wasused as injection chamber. This had the advantage of easyintroduction of sample with minimum cross contamination and issuited for automated sample addition.

For diroct estimation of uranium, the sample was added intothe cup using polypropylene tip at a steady rate and mixed withsteady flowing stream of 0.1M nitric acid with the help of amagnetic stirring bar and the mixed solution was passed throughthe photometer. A minimum but constant liquid level wasmaintained in the injection cup by adjusting the T at thedischarge point of the tubing to the same level. The maindrawback with this system was that the tubing diameter should belarge enough to allow the free flow of the solution at the samerate at which it was being delivered in to the cup. Further,complete mixing of sample and acid should take place in the cupitself.

For uranium assay by alcoholic thiocynate method using thestopped flo'w technique, the sample was added to a fixed volumeof the reagent in the open cup and after mixing, the entiresolution was drained through the photometer for absorbancemeasurement.

As an extension of this method, following the closed loopprocedure, the solution can be recirculated between the cup andthe photometer using a peristaltic pump till constancy in theoptical density is achieved and after which, the solution can bedrained off. These two techniques will be useful for allspectrophotometric procedures that require some tine for colourdevelopment. In the present work, only the stopped flow techniquewas tested.

(d) Detector Modules

For absorbance measurements a photometer (Metrohm Model 662)was used. It was equipped with a flexible, sheathed optical fibrelight guide as detector. Normally this device is to be dipped insolution for absorbance measurement and is used mainly for endpoint detection in titrimetric analysis. However in the presentinstance, a pyrex glass tube of 2 mm inner dla was inserted inthe light path of the detector tip and the solution to bemonitored was passed through this tube. Thus the photometriodetector never came in direct contact with corossive andradioactive solution. Absorbance Measurements could be made atany desired wavelength in the visible range. After initialisingthe unit to sero with blank reagent, the absorbartee of thesample-reagent mixture flowing through the tube was

- 161 -

monitored as digital output and the same could be plotted using achart recorder ( Metrohm Model E536 Potentiograph).

For gross gamma counting a 7.5 cm(3") Nal well type detector(ECIL, India) was used and the sample-acid or reagent mixture waspassed through a coiled tubing inside the well. The counting wasdone for a fixed time starting from sample introduction. Thegross gamma count rates obtained were subtracted for thebackground obtained by passing the blank solution for the sametime interval.

IV. RESULTS AND DISCUSSION

Except in the case of uranium (VI) estimation by alcoholicthiocynate method ( where sample and reagent were mixed understatic condition), in all other cases, both reagent as well asrample were introduced at selected flow rates. The colorimetricresponse of the FIA system as a function of sample and reagentflow rate was studied in detail to arrive tvi optimum conditionsas it is an important factor in FIA analysis. For this study,estimation o* ••-•»•«•• nm bv direct, colori™*"' w«a used as are/erence mecnoa.

The variations xu v-~.~ ...:ic response or the system «d afunction of reagent flowrate at constant sample flowrate and viceversa are shown in Fig: 2a, b & c. These data were useful inoptimizing the sample and reagent flow rates while standardisingcolorimetric procedures.

Fig.2 a shows the variation in absorbance as a function ofaliquot size for a given sample and reagent flow rate. It is seenthat for aliquot sizes above 0.5ml, a constant response isobserved at the detector, irrespective of the aliqout size. Thusabove this minimum aliquot size, the FIA response becomesindependent of the aliquot size aa long as the flow rates arekept constant. This is of groat advantage while planning onlineanalytical procedures that require dilution or reagent additionbefore colour measurement*.

Table I-A shows the results obtained for direct estimationof uranium in the range 40-350 g/1 using closed chamber and opencup as sample injection assemblies. As the aliquot size was0.5ml, a steady response for a reasonable length of time could beobserved at the detector. It is seen that the precision observedfor open cup injection is better than that obtained using closedinjection chamber. This may be due to the introduction of samplethrough the septum in the later case which is not as reproducibleas sample delivery in open cup. As the solution in this case isflowing under slight pressure it was possible to pass this samplereagent mixture through gamma counter (well-type). The crossgamma count rate obtained could be correlated to itsconcentration of uranium with a precision of +/-5X and theseresults are included in Table I-A. Similar colorimetric resultsobtained in the case of U(IV) estimation are given in Table I-B.In this case the stability of uranous nitrate was ensured using

- 162 -

1.0M nitric acid and 0.1M hydrazine mixture as diluting reagent.In the case of U(VI) and O(IV) precisions better than +/- 1.5%were obtained at all concentrations.

Table I-A also includes the results obtained for twoconcentrations of uranium when the samples were injected into theclosed injection chamber via permanently connected tubing(Figlc).The precision obtained in this case indicates that this mode canbe used for on -line monitoring tasks where gradual variation orsteady state concentration of uranium is to be monitored. Sampleintroduction using a six way valve is currently underinvestigation for the same purpose.

Results obtained for the estimation of uranium in the range0.5 to 4 mg per aliquot by alcoholic thiocynate method usingstopped flow technique gave a precision varying from +/~ 1.5% to+/- 0.25% corresponding to the lower and upper limits of theuranium range tested. This precision is satisfactory for processcontrol analytical applications. As aliquot sizes were small inthese cases (0.05 to 0.2ml), these could be varied depending onthe concentration of uranium without causing any major error.Thusthis technique is useful for a variety of analytical methodsthat require vigorous mixing before colour measurement.

As mentioned earlier, the same results can be obtainedusing closed loop flow technique. In this later case, a six wayvalve may be more useful for fixed volume addition, if crosscontamination is not a major factor.

V. ACKNOWLEDGEMENTS

The authors wish to express their sincere thanks to ShriA.N. Prasad, Director, Reprocessing Group and Shri M.K. Rao,Head. Fuel Reprocessing Division for their keen Interest in thework.

VI. REFERENCES

1) Ruzica J. And Hansen E.H. . Anal. Cheat. Acta Zfi.146 (1975)

2) R.T. Chltnis et al. BARC-430 BARC (1969)

3) Paul J. Worsfold, Chem. in Britain,24.1215 (1988)

4) A. Ramanujam et al, CT-25.DA1 Symp. on Radiochemistry andRadiation Chemistry, IIT, Kanpur,(1985).

Table No. I-A & B

ESTIMATION OF U(VI) AND U(IV) BY FIA

Flow rate : HNO3 = 6nl/min Sample : 1 ml/minSample aliquot 0.5ml fixed

A. DIRECT COLORIMETRY OF U(VI) AT 410nm

Sample

U«/l

350.0

175.0

116.7

87.5

43.8

350.0 *

175.0 *

Closed-cup

O.D

0.619

0.310

0.216

0.164

0.083

0.625

0.338

% RSD

0.62

0.87

1.58

0.53

1.07

0.45

0.83

Open-cup

O.D

0.589

0.306

0.210

0.150

0.083

-

-

XR.S.D

0.51

0.18

0.54

0.63

0.62

-

-

Closed-cup

Gamma counts100 sec.

3726

2046

1410

1037

558

- •

* Saaple line directly connected to closed cup

B. DIRECT COLORIHITRY OF U(IV) AT 650n«

Closed cup

O.D % R.S.D

Open cup

O.D XR.S.DU(IV) U(IV)f/1

118.0

59.0

39.5

29.5

14.8

0.766

0.432

0.300

0J229

0.125

0.60

1.10

1.50

0.66

0.68

104.0

52.0

34.7

26.0

13.0

0.705

0.356

0.262

0.214

0.111

0.30

0.33

0.29

0.57

0.10

- 164 -

SCHEMATIC DIAGRAM FOR FIA

SAMPLE -INJECTION

MAGNETIC 0AR

REAGENT FLOW

DETECTOR

PINCHCH6ITALDISPLAY

ORRECORDER

(a) OPEN CUP SYSTEM

TUBE

REAOCNT rvom

DNMTAL OMPLAYM&vunw*

a—ocncTow

( t ) CLOSED CUP SYSTEM WITH SEPTUM

MfAtfNT POJHWAT

FROM MMPLC OOWMAT—*•

OMtTM. OUTLAYOR HCCMOCM

OCTCCTOI

(c) CLOSED CUP SYSTEM (WITH OUT SEPTUM)ON LINE

FIG - a

- 165 -

EFFECT OF SAMPLE SIZE ATCONSTANT REAGENT FLOWCHAUT SHED :• I S M / M *

WAVE LEMTM:- 410 *•>

SAMPLE SKCD- Iml/m*

*C«KMT S#H0 • « • ! / • ( •SAMPLE MOt ' l UIVII

EFFECT OF VARIAtLE REAGENTFLOW AT CONSTANT SAMPLE SIZEAND CONSTANT SAMPLE FLOW

ALMUOT (BOOA) CONSTAKT '• M W l t X T M / l U(VI|I t N K I FLOW MTCV-I«l/Ml«

MMCNT PLOW RATE VMIMCO.A> IM/tlhl• * 4«l/ •*««C * tmi/mit,

D •

EFFECT OF VARIABLE SAMPLERATE AT CONSTANT REAGENTFLOWMJOUOT (SOO.UW M n . t > l 7 S « / l U{V1)

KCMENT PLOW HATE-• • ( 'mi l l•AMP1E PLOW MTC VAMCO. 'A — I mi/ ml*

• m

C «

0 . M * Ac as*

0.478

0.300

•ooJk aooxTIME ELAMCO

400JL sooA tml

FIG. 2

- 166 -

STANDARDIZATION OP A D.C. ARC CARRIER - DISTILLATION PROCEDURE ON ADIRECT READING SPECTROMETER FOR THE DETERMINATION OP B, Cd ETC., IN

NUCLEAR CRADE URANIUM.

S.S. Biswas, P.S. Murty, S.M. MaraChe, A. Sethumadhavan, V.S. Dixit,R. Kaiaal and A.V. Sankaran

Direct reading optical emission spectrometers which use photo-

multiplier tubes as detectors enable rapid analysis compared to spect-

rographs in which photographic emulsions are used for recording the

spectrum. However, the use of direct reading spectrometers in con-

junction with d.c. arc excitation source, is limited since it is not

possible to measure simultaneously the intensity of a line and its

adjacent background with these spectrometers. Due to the high b.g.

associated with d.c. arc spectra it is essential to apply b.g. correct-

ion to obtain net line intensities. Since dynamic b.g. correction is

not possible different approaches are adopted by different workers to

correct for the b.g. After several experiments, we found that 'blank

subtraction method' works well to give b.g. corrected intensitica. In

this paper we present details of a d.c. arc carrier-distillation

procedure standardised on a Jobin-Yvon model, JY-48 direct reading

polychromator, for the determination of eleven trace impurities such

as B, Cd etc., in refractory U,0_ which is obtained after igniting

uranium metal.

KEY WORDS: Ur.inium Impurities, Carrier-Distillation procedure.

INTRODUCTION:

To fulfill the objectives of DAE, to produce 10,000 MWc power

by the end of this century mass production of nuclear grade uranium

has been planned from indigeneoua resources. This called for the

development of rapid analytical techniques', related to fuel-grade

uranium technology.

- 167 -

In earl/ I960'8, a d-c arc carrier distillation method for

determining trace impurities in Uranium has been developed by the

Spectroscopy- Division. The method involved photographic emulsion

technique of intensity measurements in the determination of the

impurities. Since then this procedure has been adopted routinely

in the quality control of Uranium metal and Uranium fuel,produced by

the Uranium Metal Plant & Atomic Fuels Division respectively.

The photographic method is slow and also in recent years the

supply of photographic emulsions has been irregular. Due to this

we opted to use photoelectric method of signal detection and process-

ing and accordingly installed a JY-48 direct reading polychromator

in our Division.Changing over to P.M detection required certain

Modifications of the experimental parameters.

EXPERIMENTAL:

Preparation of standards:

Using pure VJOa a master standard was prepared such chat it

contained B, Cd at 20 ppm, Co, Cu, Mn, Pb, Sn at 200 ppm, Mg, V at

1000 ppm and Cr, Hi at 2000 ppm. High purity compounds or oxides of

these elements ware used in preparing the master standard. A set of

four standards was prepared, by successive dilution of the master

standard in order to obtain the calibration plots. The concentrations

of the trace elements in this sac are given in Table I. All the

standards were (round wich 3Z carrier-internal standard mixture which

contained 98 pares of AgCl {Carrier) and 2 parts of Ca.O, (internal

standard).

Samples.which were in Che form of uranium metal turnings were

cleaned with acetone, pickled in dilute nitric acid and finally washed

with distilled water. The samples were dried and about 1 g« of each

sample was taken in a platinum dish and heated slowly (caking car*

to see it does not catch fire) on a Bunsen burner and converted to

powder form; heating was continued for some more time till a fine

- 168 -

black powder of U_0o was obtained. Periodical crushing of the powderJ O

with a platinum spatula during heating ensured fine powder of U,0_.J 8

After conversion to U,0Q each sample was mixed with 3Z (AgCl-Gd.O,)

Jo 2 Jmixture. The AgCl contained 2Z Ca.O-.

A 'BLANK' was also prepared, consisting of U 0 used in theJ O

preparation of the callibration standards pre-mixed and ground with

3Z pure AgCl only.

PROCEDURE:

By means of the 'software' provided by Apple Il/e computer,

a 'table-list' was prepared for the analytes and the internal standard

element, whereby the corresponding element channels were activated.

By performing some initial experiments using the callibration

standards, appropriate 'attenuator1 voltages were set for the corres-

ponding channels to ensure near unity slope of the 'working curves'.

The 'delayed exposure', sub-routine in the software is opened

up. The 'pre-burn* and 'exposure' times for the individual channels

are programmed by appropriate settings of the 'start' and 'end' in

software 'conditions' sub-routine. This is indicated in Table V.

The 'BLANK', the callibration standards and the samples pre-

pared ma described above are loaded (in duplicate), then excited

under experimental parameters given in Table II. The averaged

'BLANK' intensities for various channels are stored in the Computer

Central memory. These are subtracted from the gross intensities of

the corresponding channels for the samples and the standards, inordcr

to arrive at their net intensities. The intensity ratios for caclf

anslyte is obtained with reference to the net intensity of Che

internal standard.

The averaged intensity ratios of these standards are plotted

against the concentration on double log graph, to establish the

"working curve'. The intensity ratios of the samples are Chen read

from these curves to arrive at the concentration values of their

- 169 -

respective elements^

DISCUSSION:

The d.c. arc emission spec .ographic analysis of trace impurities

in Uranium by the 'carrier-d-.stillation' technique is well established

procedure since la-. 3 years (1).

Several authors have used different excitation parameters as

well as different and varied 'carrier' compositions (2-5). Table VI

shows the various 'carrier' compositions employed by previous workers.

Hitherto the emission signal was detected by the photo-emulsion

technique. Due to the anticipated possibility of non production of

these emulsion plates,photo-electric method of signal detection and

integration was resorted to.

It involved the use of 'delayed exposure technique' developed

and incorporated in the software and certain modification of the

experimental conditions, indicated in Table IV. It also involved

modification of the for~.ila -.:eed ro compute the intensity ratio from

the conventional

Total net intensity of the Analytc

Total net intensity of the internal standard

during the entire exposure period to

Total gross intensity of the analyte - Total gross intensity ofthe analyte in 'BLANK*

Total grbss intensity of the Internal - Total gross intensity ofstandard the Internal Standard

in the 'BLANK'

during the period programmed by the 'delayed exposure' software for

each individual channels.

The above procedure was adopted since it was not possible to

apply dynamic background correction for each analyte wavelength as

well as for each exposure. The 'BLANK' correction method as described

earlier has: been adopted.

- 170 -

The d.c. arc is a thermal excitation source wi\th the sample/

standard as the anode. The emission intensity of the element

excited is dependant on the rate of volatalisation of the impurities

streaming into the arc, in preference to the matrix. Figs (1-4) show

typical examples of the volatalisation pattern obtained by the

'delayed exposure technique' for boron, chromium and manganese,Cd res-

pectively. The exposure time period for signal integration/acquisition

of individual analytes are based on these curves, explained in Table V.

The delayed exposure time of 3 seconds from the initialisation

of the d.c. arc, for all the elements was resorted to inordcr to

prevent masking of the entire spectrum due to possibility of sudden

uranium flash which may occur at times at the start of the arc. The

early shut-off marked by + in Pigs.(1-4), was used to prevent, high

background, due to carbon-burning. This resulted in improved signal/

background ratio.

The 'carrier* composition modified to 32 AgCl (containing IX

Ca 0 ) enabled producing a smooch arc, during the entire 'burn' period.

This ensured steady volatalisation of the element impurities.

The arc current brought down to 8 Amperes from conventional

10 Amperes, reduced the.arc wandering, thereby improving reproducibility.

The sample/standard charge increased from 100 mgs to 120 mgs

was in effect to increase the absolute trace element concentration on

the electrode and thus retain our earlier detection limits with

certainity.

The modification in the formula for the calculation of the

intensity ratio against the internal standard, provided correction

for the 'residuals and the'electronic noise' in analytes as well as

the internal standard channels.

The overall effect of all the above changes is the improved

reproducibility in the estimate as indicated by th* mtan standard

deviation (Table III) which hrs been calculated on the basis of 10

readings. A comparison with the earlier emulsion tachniqua is also

- 171 -

shown in t!»e last column.

The working curves (plot of log concentration vs. log intensity

ratio) shown in Pig.8 are linear having a slop nearly unity.

The limits of determination, the analytical lines used^Table

III for the element impurities are the same as in the photo-graphic

emulsion technique earlier employed in our laboratory.

The accuracy data Table VII was established on the basis of

a certified international standard, Code No. NBL-98-6 supplied by

New Brjjnswi r1: Laboratory treated in the same manner as the samples

and read on the 'working curves' drawn from synthetically prepared

standards as enumerated in Table I. B & Mn values agree, for other

elements, the certified values are below our detection limits.

Brief Description of the Delated Exposure Technique :

The 'delayed exposure1 is a software programme developed and

incorporated, whereby, different exposure timings can be selected

for each individual analyte channel after the initialisation of the

arc.

This is for cht purpose of optimum signal integration/data

acquisition.

The 'delayed exposure* consists of'pre-burn' and 'burn' sections.

The 'pre-burn' period used in the spark excitation mode > meant to

clean the sample surface (solid) prior to 'burn*. In d.c. arc mode

of excitation, this period is redundant. The 'burn' period is

straight away commenced with. This 'burn' period is further sub-

divided into delayed exposure (pre-exposure) and 'exposure' periods,

through a complex electronic circuitary.

Towards the end of the 'pre-exposure' the channels are opened

- start receiving & integrating emission signals continously till

commanded to 'end' when the channels are shut off by programmed

timings, fed earlier.

- 172 -

The entire 'burn' period can be divided into time segments

which can be arbitrarily selected. Though the 'integration' is

continous, the 'acquisation' is aade after each tine segment &

stored. In effect, therefore, the integrated intensity during each

time segment denotes the emission intensity of the analyte. A plot

of these segment intensities against time axis, shows the volatili-

sation rate curves of the analytes (Figs 1-4). These curves help

in determining the pre-exposure and exposure periods i.e. 'start'

and 'end' segments for each channel. The arrows in the Figs, are

indicative of the delayed exposure & early shut-off periods shown in

Table V.

TABLE I : Concentrations of calibration standards

Elements addedStandard No. ( Concentration in ppm )

B, Cd Co, Cu, Mn, Pb, Sn Mg, V Cr, Ni

1 0.1 1 5 10

2 0.2 2 10 20

3 0.5 5 25 50

A 1.0 10 50 100

- 173 -

TABLE II Experimental Parameters

Spectrometer

Grating type, no. of grooves

Wavelength region, order

Dispersion

Slit width

Analytical gap

Excitation source

Exposure time

Electrode assemblyLower electrode (anode)

Upper electrode (cathode)

Data acquisition andprocessing.

Jobin-Yvon Model No. JY-48 Poly-chromator with lm concave grating inPaschen-Runge mount.

holographic, 2550 gr/mm

130-415 nm, I order

0.39 nm/mm in I order

30 microns

4 mm

Stabilized d.c. arc operated at8 amp.

35 sec. TOTAL

Vdia. U.C.C. 1990 graphite elect-rode containing 120 mg of sample/standard.

3/i't dia. U.C.C. pointed graphiteelectrode.

through APPLE-IZe Computer.

TABLE III Analytical Data

Mean RSD

Concentration range < P h o t o e l e c " ? c ) (Emul,ion

plate)

Element

BCdNiMnCrCoCuMgPbVSn

249.7!)228.80231.60257.60267.70243.20324.70280.20283.30318.50317.50

0.10.110110115151

- 1.0- 1.0- 100- 10- 100- 10- 10- 50- 10- 50- 10

5.98.410.811.59.26.48.316.98.111.510.2

1091013111314-1198

- 174 -

TABLE IV Modifications of Experimental Parameters

Carrier

SAMPLE/STANDARD

Arc Current

Exposure

ConventionalPhotographic method

5Z AgCl(lZGa2O3)

100 tngs

10A

No delayed exposure

Present Photo-electricmethod

32 AgCl(2ZGa?0_)

120 mgs

8A

3.0 sees delayed expo-sure programmed in thesoftware for all channels

No early shut-off 3.0 - 7.0 sees earlyshut-oil depending uponthe element.

TABLE V Details of Pre-exposure, Exposure times & the Segmentchoice.

Element/ Pre-exposure ExposureChannel (sees) (sees)

Exposure start Exposure end(segment) (segment)

Cd, Pb

Hg

B, Cr,Mn, Co,Ga.

3/0

3.0

3.0

18.0

25.0

28.0

2

2

2

6

8

9

- 175 -

TABLE VI "Carriers" employed by some earlier workers

Sr.No. Carrier used Authors Reference

1 22 Ga O- Scribner & Mullin 1

2 5Z AgCl Dhumad et al 2

3 52 AgCl UKAEA 3

U 22 Ga2O. Artaud 4

5 52 AgCl Page et al S

TABLE VII Comparison of data for certified International Stand-ard (NBL-98-6)

Element Present Work Certified

(Values in ppm on U-metal)

B 0.22 0.20

Ni <12.0 3.8

Mn 2.2 2.0

Cr <12.0 6 .0

Co < 1.2 0 .6

Mg < 6 . 0 2 . 8

- 176 -

ACKNOWLEDGEMENT:

The authors thank Dr. V.B. Kartha, Head, Spectroscopy Division

for his interest in this work.

Thanks are also due to Shri. S.S. Bhattacharya for preparing

the Figures in the ND Computer.

REFERENCES:

1. B.F. Scribner and H.R. Mullin

J. Rea. Nat. Bur. Std. 37, 379 (1949)

2. R.K. DhuMd et. al "

Report AEEI/ANAL/25, 1963

3. UKAEA Report No. ICO-AM/S-117.,

Dept. of Cheaical Science, 1958

4. J. Artaud, C.E.A. Report 1737

1960

5. A.G. Page et. al.,

Report Mo. BARC - 862.

- 177 -

ANALYTt 3 1.0 ppn IN 5TD.

4100

1000

3600

3200

2800

2K)0

2C00

1600

1200

800

100

"/, •'••: > NET INTLGRATED INTLNSiTY, ANALYTt

ANALYIL B IN BLANK

INTERNAL STO. Go IN BLANK

^ • W ^ N E T INTEGRATED INTENSITY, INT. GTO.

- - - - - INTERNAL STO. Go IN STD.S1

1.0 26.014.0 21.C

TME !S£C(M0S>

FIC.1. UOLATILIZATIO*; CURUEC FOR aOROfi 03TAINC0 BY THt OELAYLO EXPOSURE TtCHNJOUC

30.0

- 178 -

2400

2200

2000

1800

1600

5 1400

5 1200

3 iooc

800

600

400

200

ANALYTE Cr 100.0 ppa IN STO. S1

INTECRATEO INTENSITY, ANALYTE

ANALYTE Cr IN BLANK

INTERNAL STO. Co IN BLANK

INTECRATEO INTENSITY/ INT. STO.

INTERNAL STO. Co IN STD.S4

0 3.5 1.0 10.5 14.0 11.5 21.C 21.0 28.0 31.5 ».O

TIME (SECONDS)f 16.2. VOLATILIZATION CURUES FOR OflONIUM OBTAINED BY THE OEUVED EXPOSURE TECHNIQUE

- 179 -

2400

2200

2000

1800

1600

5 1400

5 1200

5 1000

800

600

400

200

ANALYTE Mn 10.0 pp« IN STO. SI

INTEGRATED INTENSITY/ ANALYTt

ANALYTE Mn IN BLANK

INTERNAL STD. Co IS BLANK

NET INTEGRATED INTENSITY/ INT. STO.

INTERNAL STD. Go IN STD.S4

0 3.3 1.0 10.9 11.0 11.9 21.0 24.9 28.0 31.9 39.0TIME (SECONDS*

riC.3. VOUTILIZATION CURVES FOR MANGANESE 03TAINE0 BY THE OELAYt'D EXPOSURE TECHNIQUE

- 180 -

2100

2200

2000

1900

1600

1100

1200

1000

800

600

100

200

ANALYTt Cd 1.0 pp* IN STD. 51

INTEGRATED INTENSITY. ANALYTE

ANALYTE Cd IN BLANK

INTERNAL STO. Co IN BLANK

j ^ S NET INTEGRATED INTENSITY, INT. STO.

INTERNAL STO. Co IN ST0.S1

1 t1.0 28.011.0 21.0

TIME (SECONDS!FIC.1. UOUTILIZATION CURVES FOR CADMIUM OBTAINED BY THE DELAYED EXPOSURE TECHNIOUE

35.0

- 181 -

9 • 7 M 1 2 3 4 9 « ?a<HCMCENTMTIM (PPHI

4 9* ?M

FIC.8. VORKINC CURVES OF ANALYTES IN U,O,.

- 182 -

SPECTROGRAPHIC DETERMINATION OF

B, Cd AND Ni IN MAGNEST M FLUORIDE

A. Sethumadhavan, V.S. Dixit and P.S. Murcy

Spectroscopy Division

Bhabha Atoaic Research Centre

Troabay, Bombay - 400 085

An emission spectrographic Method was developed for the detei

inacion of B, Cd and Ni in Magnesium fluoride used as lining in Che

preparation of nuclear grade uraniua. The Method involves Mixing the

MgF, saaples with pure conducting graphite powder and exciting in a

d.c. arc operated at 10 A. The spectra of saaples and those of synth-

etic standards were recorded on a Hilger's large quartz spectrograph

in the wavelength region 2200-2850 8. Using B 2497.7 X, Cd 2288.0 X

and Ni 2320.0 A* lines for calibration with Ca m* internal standard,

detection limit* of 1 p.p.* each for B, Cd and 10 p.p.* for Ni were

obtained.

INTRODUCTION

Nuclear grade uraniua is produced in our Research Centre by

employing the reduction of UF, with aagnesiua metal. In this process

magnesiua fluoride is used as lining. When the final product, U is

found to be free of B, Cd (<0.1 p.p.a each), it is acceptable as fuel.

However, when U is found to contain aore than 0.1 p.p.a of B, Cd, it

becoaes necessary to analyse UF,, Mg, MgF, apart froa U. We have been

routinely using eaission spectrographic aethods for the analysis of U,

UF. and Mg (1-4). A need arose for the analysis of MgF, to determine

B, Cd and -hence we have developed a d.c. arc spectrographic Method for

estimating these two elements as well as Ni which is also often required.

EXPERIMENTAL

i) Preparation of standards

Standards were prepared using pure MgF, which was obtained by

- 183 -

dissolving pure Mg metal in electronic grade t*NO_ and precipitating

with 40Z HP. A master standard vas prepared such that it contained 200

p.p.m of B, Cd and 1000 p.p.m of Ni. A set of five standards was then

prepared using this master standard by successive dilution. The conce-

ntration of B, Cd ranged from 1 to 20 p.p.m and Ni from 5 to 100 p.p.m

in this set. All standards were ground with high purity conducting gra-

phite powder in the ratio 1:1 by weight. Gallium in the form of Ga_0_

(0.2%) was incorporated in the standards.

(ii) Preparation of samples

Fifty milligrammes of MgF. sample was ground with equal amount of

conducting graphite powder. The sample was further mixed with 0.2Z Ca.O-.

(iii) Procedure

Thirty miliigraa ss of each standard and sample (all in duplicate)

are weighed and loaded into the cavity of u.c.c 7050 graphite electrodes.3 "

Each of these electrodes is arced against a yi dia u.c.c pointed graphite

electrode under the spectrographic conditions listed in Table I.

Table I : Spectrographic parameters

Spectrograph

Wavelength region

Diaphragm

Slit width

Analytical gap

Lower electrode (anode)

Upper electrode (cathode)

Excitation source

Exposure time

Photographic emulsion

Nicrophotometer

Data processing

: Hilger's large quartz

: 2200-2850 X

: A diaphragm having an aperture 3 mm wide

used at the collimating lens

: 15 um

: A mm1"

: £ dia u.c.c 7050 graphite electrode to

contain 30 mg standard/sample3 "

-rr dia u.c.c graphite pointed electrode

d.c. arc operated at 10 A

25 seconds

Kodak SA-1, 10" X 4" plat*

Hilger's non-recording microphotom*t*rOptical densities war* converted to inte-nsities and calibration was mad* usingN-D computer

- 184 -

Results and Discussion

When MgF was directly excited in 10 A d.c. arc, the rate of vol-

atilization of D was slow and emission of Cd was low. Addition of con-

ducting graphite powder (1:1 by weight) enabled smooth burning in the

arc and increased the volatilization rate of B and also the emission int-

ensity of Cd (Fig.l and 2). The function of graphite was to provide buf-

fer action. When graphite mixed samples were excited in the d.c. arc,

the volatilization of B was completed in 25 seconds. The emission inten-

sity of Cd increased by nearly 3 times of that obtained in the case of

graphite free samples. Due to the mixing of graphite, it was possible to

restrict the exposure time to 25 seconds which helped in decreasing the

unwanted background. The reduction in exposure time from 35 seconds to

25 seconds didn't affect the lint to b.g. ratios. In the case of Ni no

specific advantage was found due to the addition of graphite. We also

tried to employ zinc oxide as a buffer. Although it served as a good buf-

fer, presence of some Cd in the pure ZnO available with us, precluded its

use.

In the wavelength region, 2200-2850 A* the most sensitive lines of

B, Cd and Ni are 2497.7 8, 2288.0 & and 2320.0 8 respectively. These

lines were free of interference from matrix Mg. These lines were, there-

fore, employed for calibration. The Ca line at 2418.7 A* was used as in-

ternal standard. The relavent analytical data.is given in Table II.

Samples were often found to contain Pe at appreciable level. B line at

2497.7 A* suffered interference from Fe line *t 2497.8 A*. In such eases

B line at 2496.8 A* was used for calibration.

Table II. Analytical data

Element Int. standard Concentration range R.S.D.Wavelength Wavelength (p.p.m) (X)

B 2497.7 8 Ga 2418.7 X I - 20 10.5

Cd 2288.0 X Ga 2416.7 A* 1 - 2 0 15.9

Mi 2320.0 % Ga 2418.7 X 1 0 - 1 0 0 10.5

- 185 -

The calibration plots (Fig.3^ were linear in the concentration

range listed in column 3, Table II. MgF_ used for preparing the stand-

ards contained a residual amount of 5 p.p.m Ni and hence the calibration

plot for Ni was made after applying this residual correction. The pre-

cision of the method was evaluated by taking 10 spectra of a standard in

which B, Cd and Ni were present at 5 p.p.m, 5 p.p.m and 25 p.p.m respe-

ctively. The intensity ratios of B, Cd and Ni w.r.t. Ga were measured

from the ten spectra and the relative standard deviation (R.S.D.) was

calculated. The R.S.D. for each element is listed in column 4, Table

II. Our method is being routinely employed in the analysis of MgF. sam-

ples received fro* Uraniua Metal Plant.

ACKNOWLEDGMENT

We with Co express our sincere thanks to Shri Shekhar Bhattacharya

of our Division for his help in preparing Che figures on Che N-D computer.

REFERENCES

1. R.K. DhuMwad, M.N. Dixie, G. Krishnaaurty, B.N. Srinivasan and

B.R. Vengsarkar ;

Report No. A.E.E.T/Anal/25, 1963.

2. S.5. Biswas, P.S. MurCy, S.M. MaraChe, A. Sethumadhavan,

V.S. Dixie, R. Kaiswl and A.V. Sankaran ;

(Paper presented at this conference).

3. P.S. Murty, S.M. MaraChe and R. Kaimal ;

Anal Lett., (7), 147 (1974).

A. P.S. Murty, N.S. Ceetha and S.M. MaraChe ;

Frcs I Anal. Cham., (314), 152 (1983).

- 186 -,

1.0r

10 IS 20TIHE (SECONDS)-

25 30 35

FIC.1. VOLATILIZATION OF B IN HgF, ; (a) WITHOUT GRAPHITE/(b) WITH GRAPHITE.

- 187 -

0.30 -

10 15TIME (SECONDS)-

20 25

FIC.2. UOLATILIZATION OF Cd IN HgF2 : (o) WITHOUT GRAPHITE-(b) WITH GRAPHITE.

- 188 -

I I I I I I I I I I I I I I I I L

B/Ga

NL/Ca

i i i i i i 1 i i

. 5 10

-CONCENTRATION (PPM)

SO 100

FIG.3. CALIBRATION PLOTS FOR B, Cd AND NC IN MgF2

ESTIMATION OF URANIUM IN LEACH LIQUORS OF

LOW IRON CONTENT - MODIFICATION OF A

SfcTTROPHOTOMETRIC METHOD t'SI.'.G U-{2 PYRIUYL, AZO) RESOHCINCL

U. Suryaprabhavathy, Leela Copal, C.S. Chowdary and Radha R.Das.

Atomic Minerals Uivlsien, Department of Atomic energy,

Begumpet, Hyderabad - 50001b.

The selective complexIng property of the neterocyclic

8Zi dye-**-(*-Pyriayl *z») resorcinol (PAR; with uranium in

presence of another complexing solution (etnylene diamine tetra

acetic acla and sodium fluoride) in berate buffer (pH 7.8)

has been employed for the rapid estimation of uranium present

in carbonate loach liquors, in low acidity leach liquors and in

ion exchange eluates, where the content of dissolved iron Is

relatively low ( £ 2.5 gin/litre ). The carbonate leach liquors

are initially treated with nitric acict to destroy the carbonate

radicals, Tht» extent of formation of the U-PAR complex (tin,

without complexing solution is 38,700 at 530 run) is reduced to

about UQjk in presence of the optimum concentrations of the

complexing solution and of the chromogenic reagent used for the

analysis, and therefore the sensitivity. However, the formation

of the uranitM - PAR is linear wltn the uranium concentration,

even in presence of the complexing solutions, as was observed in

absorption measurements at the peak wavelengths of both 530 nm

and 540 na. The calibration graph is linear for the range

2 to 20 lig of uranium per mi. when iron present is £ !>0 ug per mi

in the solution of measurements. Other metal Ions which may

t>e present In small amount* In the above samples are also masked

oy the complexing solutions. The results compare well with those

determined by.the mere sensitive fluorametric method and the

modified method enables the analysis, on a routine basis, of a

wide variety of leach liquors of uranium.

- 190 -INTK0UUCTT0N

*»-(2 Pyrldyl azo ) resorcinol (PAR; is known to be

one of the most sensitive chromogenlc reagents for uranium1-6

estimations, in the pH ranges 7 to 9, and the complex has

a molar absorptivity of 3870O at 530 nm.

The use of complexing agents like 1,2 cyclohexane

diethylene tetraacetic acid in presence of sodium fluoride

for masking the interference of several elements has been

described by Cheng' and Florence and Farraxr. In the concen-

tration ranges of the reagents used, by Florence ana Ferrar,

the tolerance of iron is limited.

This paper presents a detailea study of the use of PAR

for the estimation of uranium spectrepnotometrically in presence

of ethylene diamlne tetracetic acid and soaiuio fluoride as

ccmplexlng solution that masks significant amount of iron and

other elements that are usually present in law acidity leach

liquars where extant of interfering elements is Halted. The

method is applicable for samples containing uranium and Iran

in the proportions ol U > -JP mg/1, and Iran <£ 250 mg/1.

The interference from iron has bean further reduced by reducing

the concentration or PAR compared to the earlier reported work.

Full sensitivity of the U-PAR complex could not be retained

o . to the lormatien of the Uranyl fcDTA complex of comparable

stability; the fraction of uranium released for U-PAR formation

is a function of the concentration af EDTA ana FAR and the

accuracy and sensitivity achieved in the estimations can bt

enhanced by tha proper adjustment af their concentrations in the

measuring solutions*

- 191 -

R£AG£NTS AND FROCEDUKE:

ii) Complexing solution:- 25 gms of the dlsodlum salt «f

i£DTA ana 2.5 gms of sodium fluoride was dissolved in

200 ml of water. The pH adjusted to 8 if needed and

the solution is diluted to 1 litre.

(ii) buffer solution of pH 7.8:-

10.54 gins of boric acid and 2.87 Rms of sodium

tetraborate was dissolved in water and the solution

made upte 1 litre.

(iiij PAR*-Laboratory grace pryidyl azi» resorcinol 0.1 gut

was dissolved in water by pH adjustment to 8.0 with

NaOK and the volume made upto 100 ml.

(lv) Uranium Nitrate:- Dissolved high purity U,0. in excess

ol concentrated nitric acid, evaporated off the excess

acid and diluted appropriately so that the solution

contains 200 mg/1 of uranium and the final acidity is

at out 0.1N. The standard solutions could also De prepared

in hydrochloric , sulphuric or perchloric acids.

(v; The Proc#dure:-

An aliquot containing uranium in the range 'So ug to

250 ug is transferred to a 25 ml volumetric flask, £rom the

sample solution whose initial ptl is in th* range ox 1 to 1.5.

/tad 2 ml of the complex Ing solution, 15 ml of the borate buffer

and 1 ml of the u.1% solution of FAR. Mix after each addition

of the reagents ana make up the volume. The formation of the

cemplex is complete in 5 minutes. Measure the absorbance in a

1 cm cell against the reagent blank under similar proportion*

at 540 nm. A Varian UV-Visible Spoctrophetometer was used

for tne measurements. It is recommended that the measurement

- 192 -

are made in one hour to minimize interference from elements

caused; when present irt larger amounts in the sample. The

amount of uranium present in the sample is evaluated by

comparison with a standard curve.

RESULTS AICD DISCUSSION

A pre requbite for iron not to interfere with PAR is

that it should be completely complexed with another non-interterin|

complexing agent. Initial experiments showed that £DTA is a

strong complexing agent lor masking many metal ions usually

associated with leach liquors of uranium. It was also observed

that unlike with LCDTA the .:ull sensitivity o/ the reagent is

not achievec in the formation of the U-FAR complex in presence-

of tDXA although the extent or formation of U-PAR was linear

with, concentration of uranium in the (range 2 to 20 xig/mlef M)

when measured at wavelength* 530 nra and 5*»0 nm for a constant

initial concentration ox" Zl/TA «See Table I). Studies on the

variation of the optical density as a funcxien of diiferent

concentration of EUTA for a given value of uranium and FAR

showea that after a sudden drop of 0.0. initially, for EDTA

in the range of 0,003 to 0,006 M the change is gradual, The

ratio of the formation constants ( £2- ) for the two equilibriumK2

reactions

U • PAR K1 ^ U-PAR

U • fcETA K2 „ U-EDTA

has a value of 20 • 2.0 at room temperature. The values

determined for different conditions are summarized in Table II,

- 193 -

This value indicates that for a formation of > 9596 of the

U-PAR complex in presence of JiDTA as the masking agent, the

ratio of the complex ing agent to PAR has to be maintained as

l_ 1. however, for an efficient masking of iron and other

elements in solution it was essential that this ratio should

oe ~>i 15. A concentration ran.-e of u.006 to 0.008 K of

oUTA can be chosen for final measurements when iron is present

l_ 50 mg/litre in the measuring solution and the concentration

of PAR fixed alf 2x10~H accordingly and compared with the

standards solutions of jiranium. The value of uranium determined

in different leach liquors under different combinations of

concentrations of £DTA and PAH are summarized in Table III,

For the concentration ranges given the values obtained vary

with in 10% and are found to De in good agreement with those

determined flucrimetrically, both for the leach liquors ana

for synthetic solutions containing iron. This equilibrium also

inCicat£& that/carbonate* leach liquors where interferences sre

minimum, the sensitivity of determination can be improved by

decreasing the amount of cduplexing solution used.

A point ef interest is that CDTA is Known to form metal

complexes which are generally more stable thaH those witn ECTA.

The retention of full sensitivity lor uranium with PAR in

presence ol CDTA and reduction^sensitivity in presence of £DTA

suggest that uranyl ions apparently reacts weakly with DCTA,

compared to KOTA. The higner tolerance of iron In presence of

£DTA as the masking agent in comparison to CDTA can only be

explained on the basis of the likely bond breaking of the Fe-CDTA

at higher pHs in presence PAR.

- 194 -

CONCLUSION_

The spectrephot•metric method using PAR in presence

or EDTA as complexing agent for masking the interfering

elements has been applied to the estimation of uranium in a

variety of leach liquors. The results are in'good agreement

with those obtained by fluorimetry and by spectrophotometry

after prior seperation of the uranium from interfering elements,

The method offers a rapid procedure for the analysis on a routine

basis ana is applicable to carbonate leach liquors, low acidity

leach liquors and for ion exchange eluates. These solutions are

usually associated with low content of interfering elements. The

carbonate leach liquors have to be decomposed with nitric acid

prior to analysis. The amount of EDTA used is sufficient to

mask most of the interfering elements associated with the leach

liquors ana the concentration of PAR employed selectively

releases arid complexes the uranium to the same extent as the

standards employed. The error observed is £ 10* depending on

the total content of interfering elements that consume a part

of the masking reagent. The accuracy of the results Improves

when the optical density measured of the sample approach that

of the stsnc'ara. it is also recommended that the measurements

are completed within an hour so that the eclour enhancement

(if any; with time, due to the presence of excess iron,

niobium etc Is minimized. The method has resulted in a

considerable saving of analytical time.

- 195 -

ACKNOWLEDG iSMEKT S

The authors are grateful to Shrl 3.N. Tikoo,

Head, Cnemistry Group lor constant encouragement and

interest in the work ana Shri A.C. Saraswat Director,

AMD for approval to present the paper.

- 196 -

Table I

Fermatlen ef the U-PAR complex and the opticaldensity at different conditions,

a b e

O.D

0.70

1.13

1.40

3.50

3.50

3.50

3.50

0.52

0.70

2.10

4,20

0.70

1.40

2,80

4.20

fPAR] X 10.M

2 . 0

2 . 0

2.0

2.0

1 .0

5 . 0

10.0

2.0

2 . 0

2 . 0

2.0

2 . 0

2 . 0

2 . 0

2 . 0

O.D

0

0

0

0

0

0

0

3.0

3.0

3 .0

3.0

6.0

6.0

6.0

6.o

530m

•25D

.430

,511

1.307

1.238

1.338

1.406

0.100

0.135

0.395

0.785

0.110

0.217

0.429

0.620

B 54Onm

.246

.410

.500

1.256

1.200

1.306

1.386

0.092

0.125

0.370

0.742

0.100

0.205

0.410

0.600

a) Refera ta concentratlen ef uraniua In the range1.5 ta 12 «g/l in the aeaauring aalution.

b) Refer* ta addition af 0.5 ta 5 al of V.1% solutlanaf PAR in 2> al salutlen.

c) U ta 2 al af tha i?.5* aolutlan af ZDTA added ta 25 al selutlon.

d; The Sandal aenaitivity at the eptlaiM concentratlena ot thereagenta recewnendea cerreapanda ta 0.018yug af uraniuaper oa2. '

j\j]

- 197 -Table II

The equilibrium constant calculated under differentconditions of reagents in the formation of Uranyl-

PAR Complex from U-SDTA.a b c (d)

x 10,K [EDTA] x 10,M [PARj x 10 , " K

1.41.41.4

1.A1.4

1 .4

2 . 8

2 . 0

2 . 8

2 . 8

3.53.53.53.53.53.53.53.5

3 . 0

3 . 03 . 0

6 . 0

1.5i.O3 . 0

3 . 03 .0

6 . 0

1.53 . 03.03.04.54.56.06.0

1.2

2 . 00 . 8

2 . U

1.01 . 0

1.2

2 . 0

0.82 . 0

1.0

1.0

2.55.02.55.01.0

2.5

18.318.718.022.019.522.519.518.51O.0

21.021.924.020,022.521.619.324.021.6

a) Variation of uranlua in solution i s 4 mg te 10 mg par l i tre .

b) ComplexIng solution per 25 ml la 0.5 ta 2 ml.

c) O.fc to 2.5 ml of 0.10% PAR per 25 »1 solution

a; The average of the measurements at 530 nm ana 540 run.

- 198 -

Determination of Uranium in airrerent samples. Using

different initial amount or KDTA ana PAR. Berate buffer

added is 15 ml in 25 ml solution and measurements at 540 nm.

PAR,0.1%

ml per 25 ml

1.0

1.0

0.6

0.4

1.0

1.0

0.6

1.0

1.0

1) a and b reler to two dilutions or tho originalsample used for the measurements.For SI.No, 1 ted, tho dilutions are 1250 and 325respectively and for 5 to 7 tho dilution are 100 and 50.

2) The sample of serial numbers 1 to U was analysed tocontain 12.5 ga/lltre of iron*

3) SI.No. 8 and 9 refer to a synthetic mixture of uraniumana iron; in different quantities

Q :> Uranium la 2uO mg/litre and iron 1 gm/litre.

9 - > Uranium la 200 ag/litre and iron 3 gm/litre.

The dilutions are 100 and 50 times respectively.4) The allquots of the samples were analysed by three

different analyses using different composition ofcouplexing agent and PAR and the values agreedwithin 0.5%.

•

1

* 2CO.

t>

u

5

6

7

8

9

^ Complexingsolut ion,ml in 25 ml.

1 . 0

2 . 0

1 . 0

1 . 0

1 . 0

2 . 0

1 . 0

2 . 0

2 . 0

UraniumBn/litrea3.20

3.14

3.06

3.01

0.272

0.26y

0.270

0.197

0.205

found

b

3.40

3.18

3.12

3.08

0.275

0.267

0.274

0.203

0.212

- 199 -

REFERfiKCES

1* Cheng K.L.Anal. Chem. jK>_ 1027 (1958).

2. Folland F.H. t Hanson. P ana Geary, w.j.Anal. Chim. Act*. 20_ 2b (1959).

3. Busev. A.I, and Ivanov. V.M.Vestnik Mnsk«v. Univ. Ser. Khira.N«. 3, 52 (1960).

<*. Cheng. K.L.Talanta. 2 739 (1962).

5.. Florence T.M and Farrar Y.Anal. Ch«m. 1613 (1963)*

6. tir»lc. I , P«lla.S; and Radcsemic. MKicr*. Chin. Acta. II (3 -4) , 167(1985). ( A.A. f»8_ 10U 96 (1986).)

- 200 -

Seaaion II-B

Discussions

Paper Ho. 1

V. K. Panday s What criteria la applied for choosing the

background line for applying correction?

A.B. Patwardhan t Background wave length should be free from

any Interference due to Matrix and anolytes.

Paper Wo. 2

S.K. Aggarwal i Would you like to give an idea about tHe

detection Halt of B In U? What la tbt aeaory effect In ICP-M3?

•

T.R. Mahalingaa t Onee we reaove the uranlua aatrix by eolrent

extraction, the detection Halt will be about 15 ppb. There Is

the peak oyerlap interference froa the strong 0 peak and the

B peak. Hence, high resolution aode has to be used to get

rid of this Interference* This results in poorer sensitivity

and detection Halt for B.

Meaory effect has been noticed only If we use solutions

of high concentration {>!*). But, with solution of 0.1* salt

concentration (which is generally used In 1CP-MS) no aeaory

effect has been noticed.

R.K. Dhuawad » Has this ICP-MS aetbod been eaployed by other

laboratories abroad? If yes, are their experiences and 'rindings'

3lmllar to yours.

- 201 -

T.R. Mahalingam : Many laboratories abroad are using ICP-MS.Our findings on sensitivity, detection limits, drift and matrixinterference tally very well with their experience reported!^in the literature.

R.X. Dhumwad i Has anybody used ICP-MS for analysing Pu samples?

T.R. Mahalingam t I am not aware of any published reports onanalysis of Pu by ICP-MS* But I have seen that ICP-MS has beenalready adopted tc glove-box operation in the Institute forTransuranium Elements (European Atomic Energy Commission) atXarisruhe, Westt .Germany. They were analysing Am in the activewaste^solutions. . *

S.M. Marathe t What is the limit set for maximum solute concentra-tion? Do you experience clogging of nebullser?

T, 3, .Mahalingam s i ihlnk that you are referring to the maximum3ample or aalt concentration. Generally a 0.1* solution iseasily handled. "Matrix interferences are more at higherconcentrations.

We did not expirienoe any clogging problem even with 1*solution of sodium nitrate.* Published literature indicates thateven with refractories, unwrnxxcixm no clogging problem wasobserved, when the sample concentration was kept at 0.1* or leas.

K. Syamaundar i What la the aample sise of uranium taken for therare earth determination? What la the. detection limit for Gdfen achieved with that sample slse? What la the throughput ofsamples for analysis?

- 202 -

T.R. MAHALINGAM : The detection linit for Gd is 1.8 ppm.But once the uranium is removed by solvent extraction, thedetection limit has been found to improve to 0.002 ppm.About 5 samples could be analysed for ten elements in aboutone hour.

Paper Ho« 3

N. MAHADEVAN ; Why do you want a better or a superiorspectrophotometer for simultaneous analysis of TJVT and andUIV in your PIA system?

A.H. PARANJAPE : I f both UIV and UVI are to be measured fromsingle injection of sample then simultaneous measurement of

TV VT

U at 65Onm and U at 420 ntn would be required. The detectorwe have used is for absorbance measurement at single wave length*Simultaneous measurement would require a stopped flow techniquecombined with a spectrometer capable of automated scanning attwo wave length.

S.K. AGARWAL t Can the flow Injection analysis technique be usedIV VIfor determining the per centage of U and U ?

A.H. PARANJAPE i Y«e, TJVI does not interfere at 650nm where U I V

is measured. So it can be analysed without difficulty. IfVT • I V

U is to be measured in the presence of U then correlation ofIVabsorbance at 410 nm for the presence of U is required at nee i t

Interferes at this wavelength*

- 203 -

Paper No. 4

H.C. JAIN i What is the lowest limit of B and Cd in thehigh purity uranium which is used for making master?

S.S. BISWAS t High purity U-Og used in preparing the MASTERSTANDARD for B, Cd etc. is examined by optical-emissionspectrograpbic method. If the spectrum does not show linesdue to B and Cd, under normal exposure condition, it isinferred that these elements must be less than 0.05 ppm.AbOTe this level, B and Cd lines will be seen in thespectrum.

If we prepare calibration standard using such U-0Q,the calibration graph at the lower limit will not be linear -i.e. a 'toe' will be Indicated. The linearity of finalcallibration graph* for B and Cd using certified batch ofU~0g, will further confirm that both these elements must bepresent at 0.05 ppm level. However, we would prefer tocheck these values with S3-MS method.

A.B. PATWAHDHA? » If blank contains impurlt. nil this besubtracted?

S.S. BISWAS t Yes, the "blank subtraction" corrects for the"electrode blank" also i.e. for example Mg. Further, itcorrects for the residuals In the matrix and the noisecontributions from electronics circuitry. The overall effeotcan be seen by the it R3D In the table given in, the text*

S » S S I O H I I I A

MINING AND ORE BENE7ECIATION

Chairman : Shri A.O. SARA SWATDirector AMD

Reporteur* Dr. V.N. Pandey

UCIL

- 204 -

DEVELOPMENT OF niNINn A.T JMQUGUQA

By

J . L . BhasinChairman & Managing Director

Uranium Corporation of India Ltd3adugude

1. INTRODUCTION

In th» context of the powar reouiremant of the country

atonic powar essuned a considerable importance aa an altcrnativa

sourca of energy. Vith the limited raaourcaa of foasil and ydro-

alactrlc resources it assumed a greater importance. In tue *trst

ohaee of the nuclear reactors the natural uranium was taken as the

fuel. So it wee imperative to locate the uranium deposits in the

country to meet t*« requirement indigenously. The occurrence of

ur*nium minerals in the famous Singhbhui" Thrust <?elt was known since

1937. In 1950 a te-"" of Geologists w*r aeelgnerl the specific task of

closely examining th« 160 K" long Slnghbhum Thrust Salt. Th« team

after exploration located a number of uraniu* occurrences. Thn

deposit et Jedugud*1 eventually turnedout to be e m«jor one; it was

discovered in 1951. *fter th« discoveryy crw detailed prospecting

end exploratory mining was conducted by Atomic minerals Olvlsion of

Oepert-en' of Atomic Energy. After the exploratory minln?( the

depoelt wee taken for com">erci?l exploltetion.

Dedugude is the first mine in the country to produce

ur nium ore et a co"»-»ercl''l scle. The mined orr is processed In

the Mill at Daduquda and the concentrate in t -e rorn of Pagnlsiam-

Di-ijrinate is sent to Mucleer Fuel Complex, Hyderabad. Th« nine at

Oaduguda Is designed to produce 1000 tonnes of ore per day.

The mining of a uranium deposit is a multi-disciplinary

activity involving the services of Geologists, Mlnlng Engineer!,

Pnytists, Surveyors, Pechnnicsl and Electrical Engineers. The

activities of the vrlous disciplines are co-ortfinnted and put to

*n effective use rJurlng the exploitation of the mineral deposit. Apart

tha tachnir : orks, a number of other Jobs which m«y ••jrise ...

- 205 -

during the mining operations, ara assigned to various officers.

The mining is considered to ba a wasting asset. So whatever ora

is extracted should be adr?ed in the 'ore or reserves by Further

development of the ore-body. If this does not happen the mining

operations will come to a stand still after a period of time.

2. GEOLOGY

The South fast part of Singhbhum district is characterised

by a shear zone where recks have been folded and overthrust. This

zone is about 160 Kf" in length zr.a 2 - S'lC" wide an<i is commonly known

as Slnghbhum Thrust "alt. This bait commencing from Ouarapuran with

an £W trend pasaas through Kharaawan and 5*raikella from where it

takes a turn to Jsdugude *nd t*o**bani ending ncor 9ahr»gor-3. It is

in this belt that uranium minerals era found. Geologic?lly the

thrwst bait is constituted by archaen metasedimen's, such as mica,

schists, phyllitaa, quartzites and altered tuffs. The reck types

ara classified under two st nes: the older Chaib?sa stage and the

younger Chanjori stage. Th« older rocks of Chrib^sa *ticm heve been

thrust over the younger rocks of Ohanjori stage. The thrust contact

itself was severely sheared and brecciated. Uranium occurs in this

breedated zone in very finely diseemineted for*. The mineralisation

is structurally controlled, and is confined to shears. The principal

mineral of urenlum Is Ur'ninite (U^Ge,)*

* number or uranium deposit have been located in the

Singhbhu"! Thrust 9elt, the major one being at 3aduguda, <)hatin,

Narwapaher, Tur»mf?ih, Nane*upf Keruadungrl, Kanyaluka and R

First deposit of economic importance was located at Oaduguda and

it has become aifladged operating mine. About 4 K** from ^eduguda,

a small deposit at Bhatin 'as alto started producing. Two more mlnea

at Narwapaher and Turamdih h.tve bean approved by Government and the

construction at both these sites has commenced.

At Jaduguda there are two lodea separated by a horizontal

distance of about 60 <retres. The southern lo^s known as Footwall

lode extends over a strike length of 1000 m. from Cast to Weet.

- 206 -

This lode is not uniform either in distribution or concentration

of tha ore elempnt-.s, which is exores'-ed in the development oe two

or*? dhoots known ng the Castern ?nd Cnntral lodes. The Northern or

t^e hanging Wall Lode is noticeable only in the East Tor a strike

length of about 200 metres. Of these the Central Daduguda ore shoot

in the footuall lode is most interesting not only because it is the

longest *nd richest but also because it contains Copper, Nickel and

as associated economic minerals.

The radioactive mineralisation in ?aduguda is mainly

structurally controlled, the mineralisation being confined to the

shears which are parallel or sub-pamile] to the foliation of the

rocks, Chemic-1 analysis of the ur^niii? ore from Central 3-fduqucla

indicated an appreciable content of bass metal 9 primarily copper,

nickel *nd molybdenum which are recovered as bye-products. The wi

of the ore-body varies from 2 metre to about 25 m»tres. At depth the

lenses have over lapped each other because of lateral thrust. This

has given rise to an increase.in horizontal width and'reduction in

strike length.

-HHBLHUHOTltt

l*.'.'.*l ..aeblatX7*nlt« aoblttOrtbody

ieblit

GEES ipldlorlt*

fSBJB B»»t reoH

— 6»0 •

: A ctoft-itciion of (he orfbody.

-.207 -

In the central Daducude the width has increased from 6 to 8 metres

in the upper levels to as much as 25 metres at 434 ml. The strike

length in thp upper levels is as much as 830 metres and it h?s

reduced to 520 metres at 49S ml. But the overall volume of the

ore has remained "lmost the same and moreover the increase in width

has helped in mechanising the operations. The average gradient of

the ore-body is about 40°. The Geological section of the ore-body

through shaft is given in figure - 1

3. "KJOE Pr ENTRIES

The main entry to the mine is through a shaft. The shaft

is circular in shape having 5 metres finished diameter anri is

concrete lined throughout. The depth of the shaft it 640 metres

2nd it is equipped with two tower-mounted multi-rope friction

winder?. The cage winder is 280 K'j O.C. winder and t e <?kip winder

is 360 KW A.C. winder. The cage and skip are b lsnced by counter-

weights and tall ropes. Double decked cage is used for lowering

and hoisting persons and material. It is also used for hoisting

wast* rock. The skip having a payload of S tonnes is used for

hoisting the ore.

The she't is alco siuipped with pipe columns for

compressed air, water m^ins, drilling *nd drinking water and power

and control cables.

£.. WINE LAYOUT

The shaft stations are generally excavated at vertical

intervals of 65 Mtres, the last working level is at 5S5 metres.

The first prospecting level w*»s opsnsd at the ground

level. Subseouently a level at 30 metres above th» ground level

and it 50 and 100 metres below th* ground level usrs opened. Below

100 metres level the level interval is 65 metras and the main

tramming levels are at 165 ML, 230 H.f 295 H., 370 "I, 434 n,

495 n. and 5SS PI, crushino station at SCO "I and skip loading

st'-tionn at 605 ("L.

- 208 -

5. SHAFT SINKING AND OEEPENINC

The main shaft at Oaduguda is 640 metres, excavated in two

stages; the first stige was to a depth of 315 metres. The sinking

commenced in April 1964 and the shaft was commissioned in September

1968. The shaft was first sunk to 34 metres through the top soil

and weathered formations and lined, with reinforced concrete in

stages using steel shutter. The mucking wag dona manually into

buckets and these were hoisted by small hoists. Next followed the

construction of the R.C.C. he?dframe using a special Swedish slioform

technique. Sinking was resumed by installing the main sheave In the

he^dfrane itsel' at an elevation of 19.50 metres. The advantage was

chat the sinkine of the shaft and the installation of the winders

on top of the headframe were done simultaneously. The excavated

diameter of the shaft we? 6 metres and it wee concrete lined through-

out the length to a finished diameter of 5 metres. As the shaft was

in the foot hill, ther- u;s plenty of seepage water. One compressed

air operated pump was operating continuously. The drilling in the

shaft bottom was done in two halves.. The benchee differed in

elevation by about « metre thus giving two free faces for blasting.

To restrict the throw and avoid damage to sollars, ladders and pipes,

spiral pattern of drilling was followed. Tor mucking, a cactus grab

of 0.6 « 3 capacity in conjunction with two 1.5 m3 capacity buckets

was used. The pipe column* were extended on Sundays. Great cere

was taken in maintaining the vertlcallty and centre line of the shaft.

Tor ventilation two fans of 15 HP eecn in series were

Installed nenr t-e s'laft top with metal ductings of 50 cms, in

diameter, flexible terylene duct* were used below the metal ducts

to about 20 m. above the shaft bottom. The shaft lining was done

with the help of sllpform. The hydraulic pump was installed at the

upper level intt t a hydraulic J"eks uern used in an inverted position.

T*e concre*.* was suoplied from surf'ce in bo*.torn dlonerge hoppers

•»nd it W3s conveyer! t'irout; • n launder to tho Aides of the theft.

lining,the sntift was equipped t/ith buntonr, r'lili, rope guides,

- 209 -

pipe columns, power and control cables. The cage and its counter-

weight and the skip and its countsrweight ware then installed.

In the second stage, the shaft was deepened to 660 metres.

During the deepening operation, the production from the upper levels

was continued. For deepening, a Pilot Shaft of 3.S metre in diameter

at a distance of 21.5 metres awey from the win shaft was sunk fro*

295 TL to 660 "T.. The sinking uas done in the conventional method

using a greb of 0.25 « 3 capacity. Subsequently this pilot shaft was

used as ore pass. The main levels were opened at 370, 436, 495 and

55? metTes when the pilot shaft rearhed the required depth. It was

preferred to do «=ufflci«snt develooment of the shaft pl^t in the first

instance itself to avoid any damzge to the shaft fittings during the

subsequent extension of the crosscut and this would facilitate the

driving of cross cuts to the bore-body. In addition to these plats,

the crushing and t*e skip loading station? were also made at 560 T

and 605 (*!.. frcm each cf these levels, raises were driven to the

upper levol by Alimak Raise C Unbar along the centre line of the

main shaft. These raises were (hen finally enlarged to the size of

the m*in shaft. The shaft was then concrete lined by ellpform and

eouipped with buntons, rail guides end pipe columns. The same

construction equipments which were used in the first stage of shaft

sinking were also useo* in the second stsge. Ventilation was main-

tained by tuo fans in series with metal and flexible ducta. Top

of this raise wes covered with a steel plate in which two pipes were

fixed for plumbing, finally the pillar in the shaft between the two

stages was removed and this portion or the shaft was equipped with

buntona, rail guides, pipes etc. Tor the installation of rail guides

an accuracy of 5 mm V*M obtained. The shaft was ultimately eoui-

pped with longer guide, winding and balance ropes and power and

control cables. A small raise was made in the pillar between the

two stages.

6. LEVEL OCVELOPffCNT

Upto 295 ft, fie development and tramming drives were

Located in the or* boojy itself "•» 0er ae possible. This was done ..

- 21.0 -

to gat some ore during development and to reduce the waste rock,

from 370 ML downwards the tramming levels were made in the footwall

in w?3te rock. This has solved the problem of frequent drags in the

ore-body. c all these levels independent compressed air, drilling

water and drinking water pipe lines are provided. The main tramming

levels are provided with 3.5 tonne capacity Granby Cars. Ore from

the stcpes is loaded Into ,the Granby Cars via pneumatic chute*. These

Granby Car are hauled by 30 HP Oiessl Locomotives. The Granby Cars

dump t'-e ore intc the grizzley with the help of a camel beck ramp.

Drains are provided in the levels for water drainage, ""ain sumps

are provided at alternate levels, the water from the upper level

beinc carried to the sumc at the lower level via t e diamond drilling

hol».

7. DRIVE OEVELOPnENT

Normally the drives in the main levels are 2.4 m. x 2.5 m.

in section for 610 mr guage track. The drilling in tho development

drives is done by pneumatic drills and air legs. Burn-cut pattern

of drilling is the standard practice. The blasted rock is loaded by

Cimco 12 5 and EWCO 21 loaders into tipping tubs. These tubs ara

hauled by diesel locomotives for either dumping at the grizzley or

hoia^inc to surface by the cage. The ventilation in the drives is

provided by auxiliary ventilation uainc auxiliary fane and metal and

flexible ducts.

8. BAISS OeveLOPWCWT

In raieea also the drilllnq is done by jack hammers end

air legs. The normal aize of tKe raise la 2m. x 1.8 m. Depending

upon the inclination and the length of the raise the following

cathode are adopted.

8.1 Open Halae

In this the drilling is done from the platforms m«"de of

piinks erected in the raise. Tho accese to the face it provided by

rope ladders which are extended ee the) rale* advances.

- 211 -

B.2 Compartmsntal Raisaa

While naking tha raises with this Method, two compartment*

are made by timber stulls and planks. One compartment is used for

the ladder-way and pips columns ate. and the other for the disposal

of muck.

8.3 Raising by A.Umak Raise Climber

This is the main method for making tha long raises. The

raises with angles less than 40° with the horizontal cannot be made

AIR ANQ WATERSPRAYING

. ALIMAK METHOD OF RAISING

- 212 -

by this method. The Alimak Guides are available in different angles

for making tna required coeibination. In Alimak Raise a compressed

air operated platform moves on the guides. Each time the raise is

advanced, a fresh length of the guide is extended and anchored to

the rock. In each guide there are four pipes, two for compressed

air, one for drilling water and the fourth for blasting cable. Our-

ing the blasting operation, the topmost guide is protected by a header

plate. Tha fume clearance and ventilation is achieved by a mist of

comprasssa air and water.

9. STQPING

Stoping means the bulk mining of ora. A mineral deposit is

formed i .to various blocks by driving the horizontal levels and

555M.1..

VERTICAL LONGITUDINAL SECTION (F.W.) LODE

. 3

- 213 -

vertical or inclined raise at convenience.The top and bottom levels

along with the end raise form s mineral block. The extraction of

mineral locked up in thase blocks is called stoping. There are

numerous methods by which this can be achieved. The method of

stoping depends upon t*>e width and gradient of the deposit, grade of

ore, nature of the hsngwall and footwall and tKe nature of the ore

body. The method should be such that it gives maximum recovery, least

dilution, the ore can be transported easily to the main tramming

level and should ensure the safety oF persons working therein. It

sr>ould be inherently sarp and proven, it could be mechanised and if

circunst3nces demand it could be changed also, ^he minino method

should be chosen very cautiously as a uronn method once started

cannot b= chr?n-.ed so easily.

At Daduguda the following methods wer? adopted:

9.1 Shrinkage Stooes

Stopinc practice in this wine was initiated with shrin-

kage stoper. in ore blocks above the ground level horizon. Subse-

quently it was adopted in western sector in 100 H_, 165 T , and

230 n where the dips were favourable. Stope lengths varied from

eO to 90 metres on an average. The width varied from 2 retre to

5 metre but in exceptional circumstances it was taken upto 8 metre.

A stops drive of 2.40 metre x 2.20 metre in section and about 5 mtrs.

above tre Main tramming level was made following the footwall contact

of the ore bed". The chute raises from the main leval were made at

intervals of 10 metres centre to centre In the footwall with a small

cross-cut so as to meet the stope drive at the footwall end. This

offered uninterrupted tr?mming facility in the main level while

drawing the ore from the chutes. The stope drive was t'-en stripped

from the footwll to hanrjuall. ^ha loader which was u\ed in driving

tv«» ft tope drive uas used to handle t 'is muck <ilso. '-'here tho width

v/ari-^ions were pronounced, only a predetermined uniform width wat

opened in the stops drive. The chute raiser- tero then biilod to

give a slope of about 60° on all sides. The slice of a haight of

about ? •nstr r- was f»k«»n alonq ft strike. Tbn longth of the ...

- 214 -

drill holes, spaced 50 cm. apart was 2.A metre Tor narrow atopes

and 1.5 metra for wide stopes. The explosive used was 60 % special

gelatine and the consumption varied frcm O.dO to 0.60 kg. per tonne

of rock broken. Usually &2t to 46"£ of the broken ore was drawn

during stopino, leaving the remaining ore to serve as the shrunkpile

to provide the foot hold. On completion of the block, generally the

end chutes were emptied first followed by the immediate next onae

from either end. To reduce the dil-jtion some ore used to be left

unblasted on the hangwall side and this used to corns down later

scaling.

The main advantages of ti e shrink=ge stopes were th t it

was ? che-?p method of mining, brcksn ore coulri be stored, no

supports >j=re renuired. 9ut t"e gradient ^ad to be favourable to

allov t -<? ore to flow in tie final drawing operation. The practice

of shrinkage storing in this mine was fsirly successfull.

9.2 Open Stopes wit* timber supports :

stope blocks with flatter gradients and widths less

than 3.5 metre were chosen for open stcping. The development wcrk.

consisted in having "> central r?ise betwaen the lowar and the upper

levels along t"e orp-body. This also served as th« main entry to

th« stope. The chut* raises at 10 metre intervals were driven in

advance. The pillar raises for ventilation and entry were driven

to the upper level as the racft advanced. Level pillars of about

a metres vertical thickness wars kept for the protection of t*e

levels. Stoplng commenced on either side o' the central raise.

full f?e« of thH ore-body along t">e dip was advanced strike wise.

Orlll holes were spaced at 60 cm, to 90 cm. intervals with a burden

of about 60 cm. The holes were ell drilled parallel making'an angle

of 45° to 60° with the direction of dip and facing downwards to

direct the throw toward- the chute to minimise t'n damage to the

supporting timber props or chockmats. Holes 1.5 ". dten were drilled

and charged «>it:' ?0 * special gelatine, l-.-stlng was dono In

alternate v ift<3. ^ystematie timber supports with 200 mm dl« prop*

and 60 cm. ana 90 cm9. 9qu«re chockmats were proui:r?ri.

- 215 -

Rockbolts were also fixed In the Hartgwell near the face where

required. The props were fixed in rows in certain blocks and in

cases where hangwall was extremely slebby, ore pillars were left to

supplement the primary suoport. Howewar, these were irregular in

spacing and were not included in the systematic support.

9.3 Cut & Fill Stopas

Thin is currently t'-e main stoping method. This method has

made possible mining the increased width, improved recovery and the

extraction of irregular lsnses or ore. The fill materi=1 used is

deslimec? mill tailings. The hydraulic filling packs very well

against the hanging-wall. Over the yesrs gradual improvements have

been made in evolving the present prctice from the earlier system.

The timbered passes were changed to reinforced concrete passes cast

at site and then to circular mild steel plates. The ore-body wes

developed by making the drives of 2.4 m. % 2.2 m. section along the

footwall of the ore-body. Then either a footwall drive or a concrete

arch or a stope drive about 5 metre above the main drive war* made.

On en averao* the length of • block was about 90 metres. Two end

raises ware made to the upper level to act as service raises.

Manholes were excavated at every 10 metres pillars of about S mtrs.

were left on e-ioi side of t->o ore blocks. Two stop* raises at

either end of tha block were made to provide access to the block.

In the stopa the slices of about 2 metres height was taken from one

end of th* block to tha othar the maximum allowable height in the

stope wns 4.5 metres. After the removal of tha broken ora hydraulic

stowing to a height of 2 metres was done to give a clear space of

2.5 metres.

The main machine deployed for the removal of the broken ore

depended on the width of the ore-body. In the inception of Cut and

rill stopes scraper-* were used to scrape the muck into the chutes.

Later track mounted loaders in conjunction with tipplnn tub* were

Introduced. In wide stopes Cat/o 310 loaders ware used for tha

muckinrj operation.

- 216 -

In the present system the ore body is developed by a drive

along the footuall. Once it is developed Tor a sufficient length,

a footwall drive is made in waste rock. The distance between the

ore drive and the footwall drive depends upon the gradient of the

orebody. When the gradient is gentler, cross-cuts are driven bet-

ween the two drives and finger raises are made to act as the trans-

fer passes. Two end raises are made at either end of the ore block

which varies from 100 to 120 «i. in length. These raises are either

made by Altmak Raise Climber or manually. From the footwall drive

the ore transfer passes are made at an angle of about 55°. An

access from tha stope to these transfer passes is obtained by driv-

ing the cross-cuts. Th« planning of the footwall drive, the ore

transfer passes and the end raises is done before-hand so as to

keep the excavation and the stripping of the waste rock in cross-

cuts to th» Minimi*. A typical layout of the present cut and fill

system is given in the figure—4. The slices of 2 m. height are

taken horizontally.

i(*|t

4 :s Qii-wuMIII flop* at JUufuda mine

- 217 -

The drilling is dona by pneumatic rock drills with jack legs standing

on the muck pile. In wide stopes, stope wagon is used for drilling

uppers at an angle of 65 to the horizontal. The advantage of

drilling uppers Is that drilling can be done in advance independently.

The main advantage of the footwall ore transfer pass system

i<= that they are not affected by the drags In the footwall of the

orebody which used to happen with the transfer passes in the orebody.

Another rosin 3dv^ntage i^ thst t^ree sides being rock, there is not

m.jch we-r and tear with the result tl-at the leakage of tailing and

S:nri has been completely avoided. The mucking is carried cut by

0.76 m3 L-iOs and 310 Cavoe. On an average, aboi'f 5,000 tonnes of

broken uf? hc!s been produced fre«r a stope Dy deoloying either of

these machines. In t*8 lower levels in the western stopes, t-ne width

of the orabody '"»as Increased fro>" 12 to 30 *. These stooes are

worked transversely from Cbotwall to hanguall. In these stopes rib

pillars of 5 m. wid*r< are Ifif*: to seoarate the oanels.

10. 5T0PC riLLINC

The mill tailings are separated Into slimes and coarse

sand by Hydrocyclones. T >e slimes ere ou*pad to the telling dan

anr t*e coarse sand consisting of 60 ""• solids and 40 * watar is

pumped to tha mins vi* tirea bore holas of 75.7 '-** dla,drilled

Inclined at an angle of 45° to the horizontal. In the mine the

tailing sand is tapped from the bottom of these bora holes and

taken to the respective scopes. The hydraulic pressure is broken at

a.ich levol and the sand from one lev<*l to the next in delivered by

diamond drill '•oles. Advantage i« taken of tha hangw»?ll lode bv

drilling vertical holes from the nsnpwall lode of the upper level to

tho footw*]l lode of the lower level, rig.5 shows the arrangement at the

bottom of the hole.

The filling in the stope is done to a heiqht of

? metre ioavinq a gap of 2.1 metres from b»ck. Trie stowing

rjrade lines ore qivetn Ln t>*>p stones for uniform stowing specially ..

TAILINGS FROM MiLL

MS-Jt&MJi-fMS /'£= /i'/'- 'y<= ^•*<"'At*'"^

"75"n«n D«A. 8ORE HOLE

«Omm OIA. HOP PIPE

F«.5. SAND STOWING BORE HOLE

in ulcfa atopat. *long tha mi>nw«vs two parforatad 75 iwr C . I . pi pus

ara flxad «nd covarad by hasaain cloth. The stoulnq is ganarally

co^T-ancad from th<* b«rrlc*da and and tha u"t*r la saonr^ted both

by r i l t e ra t lon and dacantatlon.

- 219 -

11. BLASTING

Tor blastirr- in development faces BO ^ special gelatine

explosive is used and in stopes Hectorite rnd ANFO ere used. Blasting

in the mine is carried out in between the shift*, before blasting

all the holes are cleaned by blow pipes. The blasting circuit is

checked by Ohmmeter. Rhino-200 and Conswigear-200 exp?.oders are used

for blasting. The craw of one blaster and one or two helpers blasts

one or tuo fsces depending upon the circumstances, ^or more faces

extra helpers are given to carry the explosives to the respective

places.

12. VCNTIIATION SYSTEM

The m?in shaft at 3adugud«3 acts as the down cast shaft.

Two-fans'of 100.h.p.. each are installed at the mouths of the tuo

edits which were made during the exploraticn stage. Theae fans ace

PV 160/8 - TV Axial Flov fan* with two stages in aariea. 0n«? fan is

installed on the eastern aide and the other on the western side. Water

gauge developed by these fens varies from 35 to 40 mm. The quantity

delivered by each fan varies from 2500 to 3000 m /<nin.

Earlier the complete air was taken down to the bottom most

level and via the drives, raises 2nd stopes it moved up and finally

discharged to the atmosphere. This is the standard ventilation

system in metal mines. But in a Uranium nine, the redcn and its

daughters for the bottom level are carried to the top and adding up

as the air moves upwards. Recently the ventilation survey of

Daduguda dine use conducted by Central lining Research Station,

Ohenbad. The Health Physics Unit at Oaduguda wee associated with

the survey to determine the rate of radon emission from the recks

in underground. From this survey the radon and its daughters

emitted frcm each working level and the nucntity of fresh air

required to reduce t^om below the threshold limits were determined.

It may be mentioned here thnt it was not only the quentity but

nuolity of air thnt xattered. A naw ventilation system In designed

for Daduguda. Fresh air is supplied at each work inc. level end then ••

- 220 -

after ventilating tne 8topes joins fie main return. To control the

ouantity of air in aach level, regulators will be Installed in the

return air-ways. The proposed plan is shown In Fig. 6

-50ML

606 ML

II - STOPPIN6

H — RfdULATORV - TJOORTO - TOP PIMVl

FlC,SCHCUATIC PIAGBAM OF PROPOSEDVT WtLTWOBK OF jAT»Haiit>A

- 221 -

Portable fans with their suction and delivery ducts are also

installed in the wide stopes to provide fresh air near the working

place.

The ventilation in development faces is achieved by auxi-

liary ventilation, for these both centrifugal and axial flow fans

are used in conjunction with ventilation ducts of 50 cm and 30 cm dia.

As soon as the drive has reached the end of the block, a ventilation

or service raise li "iade to the upper level and regular ventilation

is established ucto that point.

13. GR'OC CQf.'TRCL

In D^duguda l"ine the lodes cannot be distinguished easily

by their physical characteristics. The rocks appear alike whether

they are ore or waste. The uranium mineral content in the lodes

is also poor. The <ninerM bpinc radioactive emits gamma radiation

which can easily be detected by electronic instruments such as Geiger

anc" Scinfillation counters. These gamma radiations are actually

emitted by the daughter products of uranium. Vhen t^ese daughter

products are in sQuilibriui* with tke parent, the ore is said to be

in equilibrium and the measurement on gamma radiation qives the

uranium eouivalent value (UjOg). So long as this equilibrium is

not disturbed by nature or by artificial means the radiometric measu-

rements are quite reliable. Jaducuda ore is an equilibrium ore.

The Geiger counter and scintillation counter have been

suitably modified to meet the rugged working conditions in the mine.

The Geicer counter is used In the form of e directional probe. A

semicylindrical lead shield of 3 cm thickness covers the probe from

one side, the other side is left open to receive the radiations. The

gamma radiations emitted by the mineral Interact with Geiger Puller

tube end produce electrical pulses. The pulse rate i<< directly

proportions] to mineral content in the rock. The Counting Rato Meter

measures the pulse rate and has built in hioh voltane power supply

to energise the Geiger fuller tube. The whole system is standardised

in * U30g.

- 222 -

13.1 Coursino pi* development faces;

The developmen' fac? has to be dressed well before taking

the measurements. Tie front of the schielded probe is first covered

by lead brick «nd the back ground level of gamma radiation measured.

The readings are tak->n at 20 cm interval starting from the footwall

corner of the drive face at right angles to the dip direction. The

back ground has to be subtracted from each reading before calculating

the l^Og value. The moment cut off value is reached a mark is sut on

the face indicating the footwall contact. The entire face is scanned

in this way and at places wnere waste bands appear or hanging wall

2*00 • •

j . j • MtrMnt of ore/wt*le boundariei on ihe developmcn: face and

shoi-botes on ihe w»Hi.

in axposad. marks are givsn. Sonatinas tha orobody is quite wide

and tha HU> contact ia not axoosed in tha driva ltaalf. In such

cssas ranularly spaced sxoloratory holas arn drilled at right anglaa

tc tha foliations anr) are logged with a G.ft. datactor attnchad to a

long conduit. 'Fiq. 7 ) Tn° datsctor !• inserted in tha hola »nd

raarfinns nra tak<>n at regular intervals. Thesa woasuramants aro

- 223 -

then used to calculate the grade and thickness of the ore body at the

face and in coursing the drive as the face progresses with each blast.

In stopes, ore and waste boundaries are demarcated with the shielded

prob°s regularly. This is very effactiva in reducing the wall

dilution.

13.2 Sulk assay of ore;

The bulk assay of ore is done with 2 scintillation probes

loused in directional lead shields. The probes are so arranged that

only one c?r is Assayed at a timp. The radiation coming fro* adjoi-

ning cars are almost coitoletely cut off. • The counting is done with

sealers, etc. The whole system is popularly known as Scintillation

-rch due to historic*! reasons. These Arches are installed in all

main tramming levels near the ore.passes. Each car as it stops at

t->e '"<rch is assayed for its l^Og content, its location noted before

it is d'jTQad into ore pass. Th? grade thus obtained is used for

calculating the run of mine grade. The data are also used for

projecting th* grade of or* available for breaking in succeeding

years.

13.3 *ss«y of samples r

Tor the assay of samples th* scintillation counter is

housed in lead shield a«s*mbly leaving a small window for placing

the sample on th* counter. Th* sample is counted against a standard

source and assay valu* in t U-jOg d*t*rmin*d.

13.4 Assay channels

Till 1976 th* mine assay plwns w*r* prapsr*d by taking back

channel samples st 2 m. intervel in the drives. Now lnataad of chi-

pping and po-.'dsring th* sampl* and than essaying it, th* U3O9 vslu*

la daterained on the spot. That* values are than transferred on

assay plan giving thickness and grnda of or* body. *s th* faces

programs th* channel assay work follows and mine assay plans are

ready in a vary short tin*.

- 224 -

14. ORE HANDLING AND HOISTING

Ore from the transfer passes is loaded by pneumatically

operated chutes into 3.5 tonne capacity Granby cars. The rake con-

sisting oF three Granby cars is hauled by a 29 h.p. diesel locomotive

to the grizzly where these Granby cars automatically tip the ore by

'Camel back ramp'. These cars are washed by a jet of compressed air

and water after each dumping to remove any ore sticking- to the bottom.

The grizzly bars are spaced at 3G cm intervals. The boulders which

do not p3ss through the grizzly are broken manually.

The grizzly finger raises or different levels join the nain

ore pass. Ore f r c different levels comes to an underground crusher

at 580 T and after crushing the ore collects in an underground bin

between 560 and 605 "U. At the bottom of the bin there is an Electro

Magnetic feeder which feeds the ore to a conveyor which in turn loads

the ore into a measuring pocket of 5 tonne capacity. This measuring

pocket loads the ore into a skip o* similar capacity. The skio is

then hoisted to surface, "t surface the skia io guided by rigid

gulden and the ore is discharged into a receiving hopper. The ore

then collects in a surface bin and via * conveyor it is transported

to the mill. The capacity of the hoisting system is 90 to 100 tonnes

per hour and the skip travels at a spued of 10 m/s.

15. PUPPING «ND DRAINAGE

About 1*00 rtfl of uoter Is pumped out of the mine everyday.

Ourlng rniny se-»on this emount increase* by about 10 %. Pumping

of wster is done in four stages. The mein pumping stations are

made ot 165 "L, 295 ft, 434 n. end 555 IX. tfuHi-stage turbine

pumpb of 60 h.p, snd 120 h.p. arc used. At may be noticed thn main

pumping ntations are made at alternate levels. Tho water from the

other levels it drained to tho lou.tr levol vl« two diamond drill

holes of 75 mm dis. The strainers are fixed on top of these holes

to avoid Any clogging.

- 225 -

During the development stage dr3ins are excavated to handle

the seepage anri the sand stouino water. These drains lead the water

either to the m3j.n sump? or to the top of tho diamrnd drill holes.

The main sumps are provided with settling tanks for the

collection of sludge. The sludge from the main sumps and frcm

the settlino tanks is cleaned either by Calighar pump or wit~> th6

help of a small bucket and overhead crane. *

15. COMPRESSED AIR

Con-pressed air in the ™ine is required for drilling,

operation of loaders anc1 ror sone pneumatic ventilation Fans. Tor

supplying compressed air three Ptlas Copco °R—9 compressors of

capacity 90 m3 per minute of frse air and one Khosle Crepelle having

cac<<city of 100 m^ p e r minute of free air are installed on surface

near the shaft. The compressed air is supplied at a pressure of

7 Ug/c<n2. About 85 m per minute of free air is supplied to the

mill for the agitation c* slurry. Cut of the four compressors,

three run at a time and the fourth in kept a* a standby. During

summer condensation of uater vapour takes olace to a considerable

extent. rcr draining the uater, water separators are installed in

t*e main airline at wurfacp ?ntf in the beginning a* the main branch

line in tie plats. Small *ater seoarators are used in the air lines

evRry loeder.

17. HEALTH HAZARDS

Cne of the chief health hazards in mining uranium

is Trom rariiaticn. Thu radiation hazards in minas are classified

as internal and external. External hazards arise out of the

radiation from the orn body within the mine ahile internal

radiation arisns from the deposition of minerals inside thn

body throunh inhalation or lngestion. In mines wher«? the ore

is of lou grade, external radiation may net c<~>use harm to henlth

-226-

but the hazards due to internal radiation are more serious as the

radioactive materials deposited in the body are in intimate contact

uiithtthe body-tissues and will be irradiated continuously until

it decays or biologically removed. Takirvg food directly by hands

which msy have been contaminated, increases the ingestion hazard.

One of the principal radiological problem in uranium

mining is the hazard in the inhalation of air polluted by radon

?nd its solid decay products. Sadon is released into the mine

atmosohore uhen ore is brok«?n. In ncn-ventilated areas and

blind ends of the mine radcn may accummulate in high conceiitrations

and may fine1 its way into t^e main stream of air.

No doubt the above hazaros seem to be alarming. The

actual mining can be done without- any risk orovided safety pre-

cautions are taken in" racoon concentrations are kepi belou the

maximum permissible limits. In every uranluT mine or other

nuclear facility, it is mandatory to have a Health Physics Unit

which monitor"- the tuork places a* well as the persons engaged

In the different operations. The International Commission on

Radiological Protection Has laid out the standards of maximum

permissible doaas and concantratlons. In an uranium mine samplns

of air, dust, water, ate. ara taken at regular Intervals and

analysed and corractiva steps arc taken wherever necessary. Tha

persons engaged in mining and milling operations are also

constantly examined as to their individual radiation doses and

they are regulerly medically examined also. Ventilation

requirements of an uranium mine ara also much higher than

those of ot-er metal mines beca-jse of the necessity to dilute

radon.

- 227 -

18. FUTURE PLANS

As has been mentioned earlier, the exploitation of mineral

deposit is a wasting asset. So constant endeavour has to be made to

explore and develope tho deposit uith depth to add the new mineral

blocks for production. further the weak links in the production

cycle have to be strengthened to make the operations safer and faster

to give a steady production. The following arB a few such areas

which are either in the implementation stage or will be taken up in

the near future.

18.1 III-STAGE SHAFT SINKING

The diamond drill holas of the deeper series indicated the

continuation of the ore-body beyond a depth of 600 metres. The

present workings will sustain the production for the next nine years.

To maintain the production beyond that period, facilities will have

to be created below 555 ft..

SIMKIHO.

- 228 -

The Ill-stage shaft sinking includes the sinking of an auxiliary

shaft from 555 PtL to 900 PU. and equipping that with other infrastructure.

It will have its own Cage and Skip. The winders for working the cage

and skip will be installed at 495 (TL. The ore from Ill-stage will be

hoisted by skip and dumped into the bin at 555 TO.. The ore from this

bin will be transported in Granby Cars and hauled by diesel locomotives

for dumping into the present ore pass system for hoisting to surface

and transporting to the mill by the present infrastructure.

The work of III—stage is in quite an advance stage. The total

cost is estimated to be about fe 7 crores and it will take about 5 years

to complete.

18.2 RAISE BORER

The excavation of a raise ore from ore level to another is quite

a dangerous and time consuming operation specially when the raises are

steep. At present an Alimak Raise Climber is used for excavating the

steep raises. In future it Is proposed to procure a Raise Borer. The

method of raisa boring consists essentially of drilling a pilot hole

280 mm in diameter from the top* level to thi bottom and then reaming

it upwards to the full siza of the raisa. The set-up is shown in rig.9.

The method is vary fast and safe.

18.3 nCCHANISED DRILLING

Orilling in hard rock is the most arduous operation in mining.

At Jaduguda, the drilling in cut and fill •topes is done either by Jack

Hammers mounted on air lags standing over th* muck pile or drilling

uppers by a stope wagon.

No doubt the stope wagon has increased the productivity to a

certain extent but still there is plenty of scope for improvement. In

the latest method of cut and fill, the filling it done vary close to

- 229 -

M* F1MC Vtt.VC(l»»tAtlJ»HIIIM)

MUM

RAISE BORINO ORREAMINO PRILLWO Of PILOT HOLE Fic.9

- 230 -

the roof leaving a gap of sbout 1 metre. The drilling is done by tyre

mounted drilling gumbos. With this method the slice height can be

increased to 3 metres and the depth of the holes also 3 metres. But

with this method unless the race is cleaned completely, the drilling

cannot be started again. The system is very suitable for wide ore

bodies. It is proposed to introduce one machine and if it is success-

ful more faces can be mechanised.

18.4 ROCK SUPPORT IN UNDERGROUND WORKINGS

When an excavation is made in underground, the rock mass

gets de-stabilised. If the rock is competent and the excavation is

not very wide dressing of the loose rocks of the roof and sices is

sufficient. But when the rock is slabby with prominent slip planes

and the span is quite wide, the rock has to be supported by artificial

means. Earlier timber was used extensively to provide the support but

as it was Becoming scarce, rock bolts were introduced. The introduction

of full column grout-bolts with 20 mm Tor-Steel has improved the under-

ground conditions tremendously, further it is proposed to introduce

cable bolting whereby the entire rock mess of the back can be supported.

Studies by the rock mechanics Oaptt. of Central dining Research Station,

Ohanbad were conducted to decide the pera-meters of the bolts, size of

the pillars etc. About 5£ of the bolts are tested as the quality to

their installations regularly.

18.5 STOPE TILLING

The filling of the stopss by deslims mill tailings is an

important part of the production cycle. Constant efforts are made to

recover as much sand as possible from the mill tailings and in case of

sny break down either in the mine or mill, the sand is pumped to the

surface paddock to be subsequently used. Recently the steps have been

taken to ensure optimum operation of cyclones. Steps are taken to

ensure the maximum recovery of sand.

19. PEHFDRriANCE

Tor the last 10 years the Daduguda nine has been performing

vary sat is factor i ly* Figure—10 shows the performance of the mine

during the last S years. I t may be seen that during these years,

the performance is above 85flC, which is quite achievement for an

underground nine.

200,000

2.40.0O0

z.oo.000

U

Q

O

1.60.000

1,2 0.000

8 0.000

+0.000

2.C8.7/9095 7.

'900-09

FlG.10. PERFORMANCE OF J A D U G U P A MIMES

- 232 -

ROLE Of SUPPORT SERVICES IN 3ADUGUDA WINE

PINAKI ROY, S.N.BANNERJEE, PI.N.SRINIVASAN,U.N.RflOHAKRISHNAN & S.O.KHANIi/ALKAR.

For executing any mining construction and production

system ancillary supporting services of Geology, Survey and basic

engineering like Civil, Mechanical & Electrical are required. The

exploitation of Uranium deposit at Jaduguda, required in addition

services of Physicists as an important help to delineate the ore

horizons which are not aegascopically visible to naked eye. These

supporting services have been organised in 3aduguda as sub-sections

of the mining department, each contributing its role to the total

system. This paper is descriptive of their contributions in the

commercial exploitation of 3aduguda Uranium deposit over the years.

A. Geology. Physics and Survey t Planning, group:

Once the ore-body parameters arc firted by surface geological

investigations and the decision of commercial exploitation is taken

then these supporting services pley an important role for the develop-

ment end production system. As an organisation the sections or

geology, physics and survey * planning at Deduguda are separately

constituted. However, functionally their roles are so interlinked

that the dividing lines at tines become only marginal.

(I) Geology sub-group

The Main objectives are i-

(I) Geological mapping and exploretion to augment ore-reserves

(II) Daily assistance in winning of the ore,

(ill) Keeping the records of sampling data, ore-reaerv* estimation,

essey plane, sections of the ore-body, end other data

pertaining to ore production, grade and depletion and

addition to ore-reserves, *nd

(iv) Tackling miscellaneous problems referred to the section.

- 233 -

Tha Lodaa:

UraniuM Mineralisation in 3aduguda is in the precaebrian

natasediaentary rocka. It is a structurally controllad strata bound

deposit. Satall quantities of sulphide Minerals of copper, nickel

& eolybdenuB) are also occur alongwith uraniua ora zones. Magnetite

is an accessory Mineral In the ursniua ore—zones. These Minerals are

being recovered aa by products during processing.

There are two Main lodes (ore-body)

(a) Foot-Mell Main lode, and (b) Hanging-wall lode. The

foliation strike of lodes are North-Weat- South-Cast having an averagi

dip varying between 30° to 50° towsrde North-east.

Toot-wall lode is the Main ore-body having a strike length

of about 600 Meters, and the Hanging-well lode is about 200 Metres

present only in the eastern Jaduguda. The parting between the F.W.

lode and H.e*. lode le about 7 0 - 8 0 Metres of barren rorMatlona.

The average width of the ore-body ie about 3 to 5 Metres,

with loceliaed width in certain ereae (-100 to -200 cordinate) of

15 to 25 Metres western and 7 to 10 Metres (0 to +200) eastern

sections or the Mine. Th» increaee in width in the weatern Jaduguda

is due to the low angle atrlke allp reverse feult. Thia fault plane

is Minsralised with eolybdenuM sulphide. In this region uranluM,

coppero nickel and aolybdenuM ores are rich in grade, and this

repreaenta the Main ora-ehoot of the Mine. The nickel enricheent la

poesibly due to the proxlalty of Metaeorphoaad ultrabaelc rocka

(lavas) - Talc-Chlorlta Schist, Cpldiorlte rocks on the f.e". aide.

In the eaatern ^aduguda the width of the ore-body is due

to Minor cross folds and drag folds.

Geological Mapping end exploration!

Aa tha Mine waa opened up the geologist had the opportunity

to verify the surface exploration date. A detailed geological and

structural Mapping waa, therefore, cerrled out for better under-

standing, of the controle of Mlneralleetlon. Tha atructural revealed

the hidden atructuree faults, folds etc, of the ore-body.

- 234 -

In Daduguda the geological group had been abla to prove additional

new ore—zones, and also extension of the explored ore—bodies For

example Parallel lodes - lode B & C, Faulted limb of the H.W. lode,

and some pocket type of ore—body.

Dally assistance;

The ore-body at O.O3J6 eU308 cut-off grade is demarcated in

the mine face with the assistance of physicist. The centre line

for the advancement of the face is marked. Similarly, the stope

drive and raise faces are guided in the ore—body. Exploratory

bore-holes and shot-holes are drilled and radiometrically logged

to prove the width and grade of the ore-body. This daily assistance

is of prime importance as any unwanted excavation in waste rock not

only dilutes the run of mine-ore, but also ultimately increases the

cost on processing.

Channel sampling points at 2 mtra. interval are marked

in the drive to assist the physics group to carry out the sampling

by shilded probe.

Ore-reserve estimation and evaluation during exploitation;

On the assay plan the ora body is divided into smaller

blocks having more or lass uniform width. By simple mathematical

msthods(geometric body), the average grade, width snd the volume of

the blocks is eslculstsd. The tonnsgs of ore is the product of

Volume X Specific gravity of the rock. (Sp. Gravity X Volume in Cu.

metres). The sum total of sll the blocks of the mine is ths total

rsserves of the deposit, (proved resarvea).

The raaarvas of ths mine changea continually, as the

deposit is bsing wgrked. A careful record of - ora mined out and

its grade, ore locked up in pillars (mining loss), dilution in

gr ds is maintained.

Once a year a balance is drawn to know the currant

reserves. This information is vary useful for planning ths

production target and grade of ore to be mined in future.

- 235 -

Geotechnical Studies;

The uranium bearing hoot recks and the wall rocks imediately

on the hangingwall and footwall sides are sheared. These rocks are

deformed and traces of folds, faults and joints etc., are prominent

in the minable zone. The prominent shears are filled up with molyb-

denite in the widely Mineralised zones. The main stoping method

followed is horizontal cut and fill with the deslimed mill tailings

as b ck fill material.

It has been observed that the hanging wall rocks are

competent, and the back (Roof) is fairly good and self supporting.

However the back and hangingwall are stitched by systematic grout

type roof bolting (1.50x1.50 spacing) and also chockmats are placed

as an additional reinforcement in certain zones. (

Certain zones in the mine the roof conditions tend to becoma

bad -(between -100 to -200) mine coordinates). In this zone the

problems of roof fall or slide has bsen idantified. The cause of the

fall is mainly due to the strike slip fault mineralised with molyb-

denite intersecting with the prominent joint plane(parallsl to

foliation strike but hawing dip towards south-west l.a. opposite to

tha uranium lode dip). In this region (-100 to -200 coordinates) the

stoping method is modified to room and pillar - with insitu pillars

of 4 to 5 metres width laft from tha FW upto tha Holybdanua shaar

plane, and tha room width is sbout 15-20 metres. This geotechnical

study has provided graatar level of safa working conditions in this

zone for mining.

A faw bore-boles (Nx sizs) have been drilled to collect

information on "ock Quality and to determine tha strangth of rocks

(compressive k Tensile strangth) through CURS in Ohsnbad.

Rock Quality Designation (RUO) is a quick and inexpensive index of

Rock Quality. Oaara, in 1964 propo*»6 a Qualitative indax basad -

a modified core-recovery procadura which Incorporates only sound

pieces of cora that are 100 an or greater in langth.

- 236 -

He proposed the following relationship between the RQD and the

engineering quality of rocks.

RQO %

Less than 25 %

25 - 50 %

50 - 75 %

75 - 90 %

90 - 100 %

Rock quality

Very poor

poor

Fair

Good

Excellant

This data helps the mining engineer in the designing of the

excavation and the support systems to be adopted.

The physical parometree of comprsssive tensile and shearing

stress as determined on representative core aanplas fro* ore-zona

and immediate hangingwall are «-

Compressive strength - 1200 to 1600 kga/Cm

Tensile strength - 75 to 125 kgs/Cm2

Triaxial atrangth - 1100 kga/Cai2 at 250 PSI confined

pressure to 3220 kgs/Cm at

3500 Psl-confined pressure.

Assistance In mine planning

Geologists ara aasociated with the mine planning call.

The dataa pertaining to the ore-body-shape, aiza and grada (dip of

ore body), and the geological dataa - rock type, thalr structuraa

(faults, sheara, joint plan* pattarneand fraquancy) form the baaia

for the preparation of layout drawings of drivaa, drifts, raises,

oro-transfers and alao in the daaign of the stoping method to ba

adopted.

fliacallanaoua work

(A) Slta selection and drilling of bore-holes for aand stowing, t-

Oeslimed mill tailings ara being uaad aa back fill in the

stopes, Bore-holat have bean drilled from surface to underground

for tranaporting the deslimed mill tailings (W4 100 ml, W2 165 ml

and E 2 230 " ! ) • Similarly bore-holes have baan drilled between

- 237 -

levels in underground for transporting of tha 3and slurry. These

bore-holes which have replaced tranaporation of the s«nd-slurry

through pipe—lines have proved to be sxtremely cost-effective as

continuous maintenance of pipe-line A replacement and the associated

^ a n d downtime has been more or less eliminated.

(6) Bore-holes for water drainage

The main sumps (pumping stationss) are located at 555 ml,

434 nl, 295 ml, & 165 ml in the mine. The water frow the levels in

between at 100 ml, 230 ml, 270 ml and 495 ml is being drained directly

to the sump through bore-holes drilled at suitable locations. This

practice is also found cost effective as for sand slurry transport.

(C) Cable bolting

In the region (-100 to -200 mine coordinate) at 230 ml we

have taken up drilling of bore holes for extended rock bolting

(cable bolting) to stitch the weak planar structures in the crown

pillar pillar portions or tha stops below 230 ml. As discussed

earlier, in this region the two major weak planes identified are

(i) strike slip fault mineralised with molybdenum and (ii) the

prominent strike joints having dip 30°-50° opposite to foliation dip.

A pair of bore-holes are being drilled (46 am size) at 5 metre

interval for cable bolting (16 am/19 M wire ropes grouted). One

set of boro-holes are drilled frost FW side of the drive at&30°-35°

angle to reinforce possible movement elong foliation plane, and the

second set bsing drilled from HW side of drive towards FUI side

(opposite diVefttion) at 50°-55° angle to reinforce the possible

movement along the major strike joints. This will enable to work

the a topes safely, and possibly to win some ore from ths crown

pillar. Though experimental the pattern of bolting has bean

designad primarily based on geological discontinuties.

- 238 -

11 Physica Sub-group

In Jaduguda Itine, a number of Radiometric techniques are

being used for the quantitative estimation of uranium ore grade. The

radioraetric method makes use of the physical property of uranium

namely Radioactivity, Generally all very old radioactive rocks contain-

ing primary uranium, contain the various daughter products in fixed

proportions to their parent, uranium. In such a state called secular

equilibrium, the intensity of the gamma radiations is directly

proportional to the parent viz uranium. This basic principle is

utilised for ore-body delineation, ore grade estimations and grade

control purposes.

Logging of blast holes

During the initial stages when the mine was being developed,

it was necessary to know the grads and thickness of the orebody

exposed in mine faces so that the drives might progress in ore,

thereby reducing the cutting of the waste rock, ^hg radiometric

logging of blast holes provided a Method to get an accurate idea

about the direction of advance of the drive after every blast and

ths data regarding thickness, Qrade and the average grade of the

blasted rock was made available.

The instrument set up consisted of a Geiger Duller Tube

detector enclosed in e moisture proof housing ettached to long

conduct pipe. The detector is connected to a composite count rate

meter with provision for ths supply of EHT necessary for the detec-

tor and suitable electronic circuits to sverage out ths detected

eignals. The gamms radiations emitted by the volume of rock

surrounding the detector ere converted into electrical signals and

read on the count rate meter which indiceted the intensity in terms

of Current. rhe instrument ie calibrated against known standards

before use.

The holes drilled in the mine fees ere logged,by inserting

the conduct and the intensity of the radiations in terms of % U308

are determined et diecrete depths. From these observations,

- 239 -

the average grade per hola is computed. The locations of all the

holes with reference to a rectangular co-ordinate system are noted

and the same plotted on a suitable scale along with the average grada

values of the respective holes. The average foliation dip is also

nurked in the plan. From this plan, the data regarding the following

are obtained.

i) Boundary between ore and waste

ii) Thickness of the orebody

iii) Grade of the orebody

iv) Average grade of the blasted rock.

In places whera the full width of the orezonea are not

exposed in the drive itself, logging of shot holes drilled at right

angles to the foliations on both the walla helped to give the thick-

ness of cncaaled ore in he walls. Since the volume of rock sampled

by this method is much more than that of the conventional channel

sampling, the logging data are more representative than chip sampling.

The comparison of the representative scoop samples have shown that

the average grade of the face estimated by logging agreed fairly

well and that it does not depend on any particular pattern in which

the holes are drilled on the face.

Face scanning by directional detector

The logging of blest holes was carried out prior to

charging the holes with explosivea. The logqing process took

considerable time for a face containing 30 to 32 holes of about 1.25

metres in depth. Many a time situations arose when blasting schedule

could not wait for the completion of the logqing operation. With

the tempo of the progress of the mine development and preparations

for stoping started picking up, this constraint became acute and

the blast hole logging method bad to be dropped altogether. An

alternative radiomatric technique consuming much lees time to

delineate the mineralised zones was developed uaing a directional

detector. The method consisted in scanning the mine face ...

- 240 -

with this detector placed ^n contact with the walls at regular

intervals. Since the radiations S O M in all directions in a Mine, for

a meaning-fal estimation of the grade of the face, it is necessary to

shield the detector from the radiations coming from all directions

except froa the area against which the detector is placed. This was

achieved by enclosing the detector in a seal cylindrical load shield

of 3 cas. thickness (^igure-'i). The other accessories reaained the

saae as that used for logging of blast holes. The actual practice

consisted in drawing a diagonal line at right angles to tfie direction

of foliations and measuring the detected gaaaa radiations with the

probe placed across the line at intervals of 15 centimetres (Fig.2).

The measured intensities ere converted into grade in terms of J&J30B

by calibrating the directional probe with standard source which

simulated a mine face. Froa the aeasured grade values, the boundariee

of the ore zones are delineated end the average grade and thickness

calculated. The response of the shielded directional probe depends

on the average grade of the material contained in that part of the

face against which the probe la placed and extending in depth of

25 to 30 cms. Thr values indicated by this technique are therefore

aore repreaentative of a larger voluae of rock than those given by the

conventional groove samples cut an inch or aore deep on the face or

back. Looking into the enoraoue tiae and labour Involved in cutting

chsnnel seaplee, a study MBS undertaken to explore the possibility

of replacing the conventional Method with radloaetric scanning.

The studies were aede on 22 chennele by both aethode. The overall

erithastic average, of shielded probe aessureasnts when compared,

showed sbout 1% higher vsluee than that of groove ssaple values.

These variation could be oxpleinod that the shislded probe looked

into e lsrger volume end gives an overall integrated values wfieress

the channel gsvs discrete veluee. Theee obssrvstione were eleo

repeated by putting the probe across snJ along the foliations* It

wss observed that these resdlngs agreed within statistical llalte.

On the bssis of these dsts, the conventional channel eaapling has

been completely dispensed with. Presently rsdioastrlc scanning is

being done in the psck of the development drives with a ssapllng

-WOODEN HANDLE

CABLE

LEAO BODf

SHIELDED PROBE Fia.1

MAWOHC OF OAl/WAfTI •OUNDAmtS ON THE F«>DEVELOPMINT FACE jt 3H0T HOLES OM THE WALLS

- 243 -

interval of 20 cms and channel interval of 2 metres. The individual

values are transferred on to the corresponding channel for the

preparations of the assay plan of the mine. This method is in vogua

at Jaduguda since last over ten years or so.

Grading of ore in mine cars

For economic winning of ore and grade control, it is necessary

to make a quick estimation of the grade of ore before it is sent to

the mill for subsequent processing. This will enable to eliminate

that part of the ore which is below cut off grade thereby reducing

the hoisting as well as extraction casts. Another important advantage

is that with the knowledge of the grade of run of mine ore from

various stops blocks, it is possible to blend the ore to feed on

optimum grt\de to the mill. During the year 1960, Atomic Minerals

Division developed and installed such a facility at Adit No.4 in the

ground 1BV21. An arch of 12 GH detectors provided with shielding and

collomation arrangements was fixad. The gamma radiations from the

ore contained in the tub^are detected by the counters. The resulting

signals after suitjble amplification are mixed and passed on to a

precision linear count rate metar. Tho read out of the counting

rate meter was calibrated in terms of )&J308.

Latar during 1964-65, a sacond bulk assay system was establi-

shed at Adit No.2 also in the ground level, from the experience and

tha difficulties encounterad in uaing GH count ra, the sacond aystern

was furthar improved uaing tha more efficient scintillation datactora

with special collimated lead shields to cut off the radiations

;omming fro* adjoining cara (Flgura-3). Two datactora wara fixad -

on the aidas and a third one on th« top to look Into th» antira

natarip1 in the tub. Tha detactad signjla, instead of passing to a

covintratemater, wara applied to a dacade counting system to giva

tha total counts dir. ctly. In actual practice tha operator poaitiona

the tub containing thu or* symmetrically batwasn tha datactora and

recorda the total counts for a pariod of 20 seconds. Tha background

counta of the system alao counted for 20 seconds La daductad to giva

the nat counts dus to tho sourca. Tha nat counta whan multipllad ..

CABLE

PROBE

SCINTILLATION PROBE *%3

LEAD SHJELO

- 245 -

by the calibration factor of the system gava the grnde of the ora

directly. For calibrating the system, over a period of tima 68 tubs

filled with ore of various grades coating from different locations of

the mine ware positioned und*r the detectors and the net counts of

each tub for 20 seconds was observed. • The Material in each tub was

spread over a flat surface andla representative sample was drawn.

Thi? sample was chemically assayed for its U308 content. From the

net counts and corresponding grades, the average system response

for the average grade of 68 tubs was established and was taken as the

instrument calibration factor. The response of this systam depended

on whether the tub contained uniform grade of rich or poor ore. If

tha tub contained uniform and homoganious grade, the instrument

estimates the grade accurately. If tha tuba contained good ore mixed

with some poor grada ore, the instrument predictions of the grede

may be in error. These variations had bean round to be within +20}t.

However, over a large number of such tuos, the affect due to mixing

got averaged out and the predictions of average grade was quita

accurate. Later when the higher capacity (3.5 tonnes) grandby cars

we/e introduced to accelardte production to ratad capacity, tha

gaomatry of the detector arrangements had to Jbe modified to suit tha

new conditions. The aystarn was again recalibrated by rilling tha

grandby cars with known grada ora from 3 small tuba (1.1 tonne) over

a period of time and a fresh calibration factor for the instrument

was arrived. Tha radiometric bulk aasay facilities are installed in

all tramming levels near the orepjss.

To determine the overage grade of ore supplied to tha mill,

this systam is extremely rapdi and tha response quite accurate The

system is in use aince 1968 and the entire ore grade control a, grade

assessment and projection of expected average grades for subsequent

years are boing done pased on these techniques.

- 246 -

III Survey & Planning; Suo-Group

The opening of an ore deposit Tor commercial exploitation

sets the stage fjr organising this sub—group which has an essential

service role both during nine construction activities and later during

production stage. This is particularly so for underground operations

where all the openings commence fros blind ends. At Jaduguda the

planning part is essentially a co-ordination with geology, physics

and production groups for preparing advance layouts initially for

mine openings followed by block to block design systems for winning

of ore. The survey part of the sub-group basically functions tp

translate the planned ideas and designs on ground. Unlike fixed

surface installations in Manufacturing industries, mining is a

continuous process and the design systems have to accomod4te sub-

sequent changes at production stage as more detailed data is revealed

about ore horizons on opening. An advanced perception at planning

stage is thus attempted by co-ordination and inter-action with these

sections to minimise the likely changes that may have to be confronted

with and compromised during production. Any subsequent ch.-inge in the

planned layout not only hikes the output cost but brings in hurdles

in meeting the day to day and month to month out put targets.

This sub-group has thus been engaged in this role at

Deduguda mine covering the following important functions:

1. Preparation of pra-project stage surface plans and drawings to

help the planning process in respect of surface layouts for

basic mine entry systems and for positioning ancillary surface

structures required for underground mining complex.

2. Rendering positional and alignment control assistance for surface

layouts during mine construction.

3. Survey control of excavation, size and verticality of the main

vertical shaft, which is Jaduguda nine's principal mine entry,

during sinking. Besides controlling the main shaft excavation

this work included opening of shaft plots as per designed ...

- 247 -

layouts at various depths and also all the assistance required for

the fittings and fixtures for two multi rope friction winders

together with the loading and unloading arrangements for ore winding

by skip and the landing arrangements of the cage winder in respect of

positions and alignments. Tolerances given for any Misalignment for

these installations by the designers were in fraction of milli-

metres and this, therefore, left very little elbow room than to achieva

the exacting standards.

4. Control in respect of alignments and gradients for all underground

development work through shaft plats for approach tunnels to ore

horizons, ore drives, tramming and haulage drifts, and all other

permanent excavations for electric sub-stations, resarvs station

first-aid rooms, main sump etc.,

5. Secondary survey control for stope blocks for their entry raises,

transfer passes, backfill gradients, volumetric measurements etc.

Besides the above basic functions the survey sub-group also

shoulders the statutory responsibility for maintenance of mine plans

and sections and their continuous up-dating as mining proceeds. Any

excavation made below ground should be corrslatable to features on

surface end elso amongst various openings made from horizon to horizon.

At any instant during the productive life of the mine, the relative

position of all workings are to be known precisely such that the

advancing fVcss, whether of development or stope, do not inadvertantly

meet across or hole through into the other workings without proper

warning having been given for withdrawal of men from the likely zones

before blaeting. Any mishap to life or demage to vital equipments

due to the incorrect end erroneous poeition of advancing faces shown

in the drawing*, is the statutory responsibility of the nine Surveyor

under the nines Safety Regulations. The relative positions ere

determined by precise traversing and plotted in drawings with refe-

rence to X,Y * Z co-ordinate systsm with sppropriate correlations

from surfaca reference grid and benchmark for the third dimension.

This role is tptly fulfilled at Jadugude nine ell these years end

there has been no accident of eny kind attributable to erroneoussurveys and computations.

- 248 -

Surveying instrument* so far being used at Jaduguda, Bhatln

and of late for works relating to naw projects are 20 seconds micro-

optic Theodlites for angular observations and precise tilting Levels

besides standard apring steel tape bonds for distance measurements,

fhh standards of accuracy for obtaining relative position of

workings in XfY & Z co-ordinate system has varied from 1:2000 to

1:10000 depending on type of surveys and the end use of the obser-

vational data. Where higher order of accuracy was required the

results have been achieved by repetition of observations and the

corrective processes of taking means and distributions of errors a»

per standard procedures. Thase conventional nethods are very

arduous and time consuming particularly in respect of transfer of

meridian through vertical shafts for correction of mine workings

by using plumbing systems with thin wires. It is practically impossi-

ble to bring the plumb wire suspension to true vertical!ty as 100jt

dampening of the oscillations is not achieved even when the plur'..

weights are freely suspended in high density oil medium for this

purpose. Certain deviations are, therefore, taken for granted.

Surveying systems have been considerably modarnissd, Ofay,

revolutionized with the application of Electronic Oietsnce Heaaurlng

(E 0 n) devices and uaa or Laser beame for alignment control and for

correlation surveys in conjunction with Gyrotheodolites. Moderni-

sation has also baan effected in tha sphere of survey calculations

where completely computerised softwsras ara available. Theaa

davlcaa yiald not only far aore accurate raaulta but ara also cost

affective in view of aaving on tin* loat in conventional systems.

With an aye on coat effectiveness due to rielng labour costs in UCIL

•lnas certain modernisation and updating of techniques in this

reepact is envisaged. To bagin with it is proposed to introduce

Lsssr ayapiace with optical plummet aa replacement for shaft plumb-

ing davicaa both for correlation and shaft alignments snd their use

for fittings and fixtures. Use of Laser beam for ahaft plumbing

have shown standard deviation of 0.14" to 0*16" upto a depth of 2 to

3 kilometres. Tor pracision levelling lnatruaentt uith optical ..

- 249 -

oLmechanical compensators (Automatic Level 4 precise staff) have already

bssn introduced. Introduction of parallel plate Micrometer in con-

junction with auto levels is also being thDoght of for use in levelling

base plates of sensitive Mechanical devices like winders etc.

8 ENGINEERING SERVICES GROUP

I Mechanical Sub-group

While any supporting service in an industry contributes to

its final output in one way or the other, the sphere of engineering

services definitely renaln the backbone and is the one that takes the

major brunt. Technological advances directly reflect on the equipment

that one uses and the outputs thereof. Keeping pace with its Modern

mining methods, the Mechanical engineering arena at Daduguda dines

has taken significant strides in the Mining of uraniuM. From a

humble beginning in the year 1961 with just few track Mounted low

capacity loaders, a couple of jack-hammers, pumps, a smell compressor*

Jaduguria nines to-day has an array of modern loaders, drilling jumbos,

turbine pumps, high capacity comprsssors, locomotives and one of the

most sophisticated hoisting systems. A short-foray into some of the

important areas of mining shall high-light the importance of these

equipments.

Prilling

Rock drilling happans to be the backbone in any mining

industry since blasting can only be done after a hol« is drilled end

only then can the ore be collected. However, drilling into rocks

having compressive strengths of 1200 - 1600 kg/cm is no mean task.

This is achieved by pneumatic jack-hammer drills which have per-

cussive (reciprocating and rotory) motions. Thess jack-hammers are

supported by pneumatic air lege having varying feeds. The Telsdyne

Upper Stopor is a self propelled two boom hydropneumetlc drilling

machine that can drill upwerd holee much fester. While compressed

•ir is the prime never, rest of the major movements have ere all

hydraulically controlled.

- 250 -

Loading & Tramming .

Shifting the blasted and broken ore from the stops (mining

area) to the hoisting area involves the use of loaders, tramming cars

and locomotives. With constraints of space and handling difficulties

it is imperative that these ore handling equipments not only be

compact and sturdy but also Manoeuvrable and fast. The earlier low

capacity small track mounted loaders have given way to pneumatic

tyjfed hydraulically operated loaders (or L.H.O's) as they are called.

The operators are comfortably seated on the loader when they work,

thus causing minimum fatigue. All controls are economically placed.

The L.0.0*8 are furthar supported by the pneumatic controlled Cavo,

Hoppar Loaders and 824 Loaders. The Cavo Loader is an imported

equipment while the rest are indigenously manufactured. However,

track mounted loaders are still being usad in various underground

development faces.

Aftsr tha loudara have duMpad the ora into tha ore transfer

chutes (which have pneumatic oparatad gataa) tha ore ia dumped into

tha Cranby Car which ara tippling wagons hauled by 30 H.P. dienel

locomotives, and dumped into the main ora pass.

Hoisting

Lowarlng and hoiating of nan and matarielsis by double dack

cage of 3.5 Tonna capacity and ora by a 5 Tonna capacity skip

comprises tha main hoiating systa*. ^aduguda had baen a fora-runnar

in installing a sophiaticatad system of winding known as tha Koepe

friction finding System. With epaada of 10 m/aec. for heiating of

ore a high output ia obtainable. The entire system can be put in

the automatic mode which thereby effects auto synthronised movemante

in the entire range of loading, weighing and hoiating operations.

Compressed Air

Compressed air is the virtual life-line in Jaduguda Mines ee

almost all production equipment viz. drill machine, loaders etc

operate on compressed eir. With four high capacity compreesore ..

- 251 -

having a total generating capacity of 13000 cfm it is Imperative .that

the compressors are kept, in proper running condition*

Pumping

nine water happens to ba an unavoidable irritant which has

to be pumped from depths of 640 M. This is achieved by means of

nigh head and capacity multistage turbine pumps installed at four

main underground pumping stations thus bringing about a four stage

pumping cycle. Water to the tune of 3 lakhs gala.is pumped out

every day.

Apart from the major equipments cited above, the mechanical

engineering section has a workshop, fabrication shop, blacksmithy,

carpentry and rock drill shops. All these shops essentially render

back up services to the mining equipment maintenance apart from

fabricating and supplying daily items like ladders, rock bolts,

crossings, chock-mats etc. To minimise downtime of equipment

certain critical machine rosiponento are kept as spares to enable the

damaged or broken part to ba replaced expeditiously. With equipments

working all over the mine a centralised information system has been

formulated so that timely action can ba Initiated. Strict maintenan-

ce schedules are followed for practicably all equipments and

specially for the hoisting system.these are very stringent. Condi-

tion monitoring devices are used from time to time to etudy various

facets of the equipments and lubrication surveys are also carried

out. With the advent of new equipments s greater emphasis has been

laid on the training of personnel viz Mechanics, operators etc.

Besides, an alround effort is always on towards indigensation of

sperea and equipment to reduce pressures for their import.

Corrective Maintenance end Technology upqradation

Emphasis on technology upgradetion vis corrective melnte-

naance has been a constant endeevour in the mechanical section.

Changes made in certain equipments like the imported Teledyne

•toper and the indigenous Hopper loader havo increased the efficiency

and the availability or the equipments and furthermore the changes

have been incorporated by the reepective companies in their ..

- 252 -

supplies all over. Stringent maintenance schedules vigorously

followed and non-destructive tests carried out on Shaft winding

rope has made it possible to increase the rope life from the stipulated

period thus effective saving on vital down time and costs. Use of

suitable resin coatings on balance ropes has brought about considera-

ble reduction in wear and consequently increased rope life. However,

all changes made are done keeping in view the prime concern of safety

and if any action contravens safety regulations, it is immediately

abandoned.

As the mechanised mining industry world wide tdkes giant

strides in the movement of heavier loads, deeper holes and faster

systems, a proposal for further modernization and upgradation of

equipments is also afoot at ^aduguda ("lines. The near future may

soon see raise borers, electric L.H.O's and hydraulic drills along

with the latest hoisting developments, thereby bringing about a

drastic change in the equipment chain but at the same time give

utmost emphasis towards conservation of energy and other related

cost factors.

II Electrical sub-group

With ever increasing dsmand for minerals over the last

decade the mining industry had to mechanise widely and more and more

use of electric power became a necessity.

The Jaduguda uranium mine uses electric power in almost all

spheres of mining activity and the bssic function of the electrical

section of the mine is to cater to the need with an eye to modernise,

indegenise and improve upon the facilities.

From thc> very inception the mine woe equipped with

various imported machineries particularly the winders end compressors

which were imported as packages, 'he winders had been giving conti-

nuous service for over 20 years -ind the life of various major moving

part like the (Motor Generator^ Set) of the Word Leonard System

were coming to an end. fhe cago winder i.e. the man winding system ..

- 253 -

has been recently revamped from the original Ward Leonard System to

a complete thyriato-rised system capeable of operating at a higher

efficiency level. This was taken up as a modernisation project.

The original compressors were imported Atlas Copco compress-

ors with ASEA, Sweden make starting gear. There were Air Circuit

Breakers for switching on the main 6.6 KV power to the prioiB movers.

These switches needed replacment. A survey of the indegenous market

was done and the said A.G.B's were replaced by vacuum contactors of

indigenous make quite successfully. The necessary circuit and

counting modifications were also carried out here. Further for the

compressors the motor generator sets for feeding d.c. power to the

rotor of the synchronous motors have been replaced with higher

efficiency rectifier system.

At present the sine is being deepened to about 900 mtrs.

The general proposal is to have a auxilliary shaft and complete

accessories and fitments. The electrical section is also involved

in planning the necessary power requirements end the distribution

system. In future the execution of these shall also be carried

out by the section itself.

III. Civil Sub-group

This sub-group is engaged in construction of R.C.C.

support systems wherever required in nine, construction of trans-

fer passes in stopes, foundations for major equipments, lining of

shaft walls etc.

Acknowledgement

Our thanks are due to Shri J.L.Bhasin, Chairman and

Managing Oirector, Uranium Corporation of Indie Ltd, and

Shri M.K.Betrs, Advisor, for their encouragement to present •

paper on "Role of Supporting Services in Jsdugude Mine".

- 254 -

RECOVERY OF URANIUM CONCENTRATEFROM COPPER TAILINGS.

• •• *•»

S.CHAKRABOPTY, U.K.TEWARI & K.K.BERI

Association of uranium mineral with copper ore ofSinghbhum Thrust Belt was known for quite long time andefforts were made to recover them economically fromtime to time.

The first attempt to recover uranium from the coppertailings was made in the Moubhandar works of M/s HCL/ICC( the then *Indian Copper Corporation* around the fiftiesand sixties. The experiments met with little successmainly because of the data available in this field wasvery meagre* the shaking tables used were of primitivedesign with very poor efficiency* Jigging, flotation,tabling etc. were also tried, but in vain. The projectwas subsequently abandoned. With the opening of theSurda mines in the early sixties, which reported higheruranium values of around 0.01 % U308 and also a treme-ndous improvement in the physical beneficiation techno-logy, the work on separation of uranium from the coppertailings of the Moubhandar works of M/s HCL/ICC was

* Deputy.Supdt(Chem)

•• Addl,Supdt(Chem)

••• Chief Mill SuperintendentUranium Corpn of India LtdP.Os Jaduguda MinesSinghbhum, Bihar,PIN 832 102.

- 255 -

again taken up by the Bhabha Atomic Research Centre,

Bombay and the Uranium Corporation of India Ltd., based

on these tests a full scale plant utilising iaprovod

wet concentrating tables Deister Diagonal suitable for

coarse as well as sliay particles, as main physical

benefication equipment with a capacity of 400 MT per

day was commissioned in early 1975 at Surda utilising

copper tailings from South Bank Treatment Plant concen-

trator of M/s HCL/ICC,

A typical minerological composition of the copper

ore at is :

Quartz : 62.3% Chlorite t 22.3%

Apatite t 2.3% Touxaaline * 3.6%

Magnetite : 3*2% Other transparentminerals s 0,4%

Other apaqueoxides t 0.1% Sulphides s 5.8%

The ore contains around 0*01% U308.

By the time Surda Uranium Recovery Plant was ooani~

ssioned with 24 nos. of tables, the South Bank Treat-

ment Plant of M/s HCL/ICC treating copper ore from

their Surda Mines, expanded their capacity to 1000 MT/

day from 400 MT/day. Before the taking up the expansion

of the Surda Uranium Recovery Plant, the Corporation

imported 1 no of Reichert Double Cone Concentrator with

the necessary accessories eg. hydrocyclone pumps etc.

This equipment was said to be suitable for separating

high density particles and does not have any moving

part, and had a capacity of 800 - 1000 MTPD/unit. Our

presumption was at that time to treat entire tailings

- 256 -

on the equipment as a preconcentrator and upgrading the

concentrate obtained from this equipment on wet concen-

trating table already provided during the first stage

of the plant, thus reducing the total nos. of tables to

a great extent.

Main features of ROCC :

1. High capacity

2. Low installation cost

3. Low operating cost

4. Superior metallurgical performance

OPERATIN3 PRINCIPLES OF BDCC *

The Reichert cone concentrator is a flowing film concen-

trator related to the pinched sluice concentrators.

High specific gravity particles are concentrated to the

bottom of the flowing fila which comprises a suspension

of solids in water with a normal solids s water ratio of

60*40 by weight*

The separation mechanism is the gravitational hindred

settling and interstitial trickling of the high specific

gravity and fine particles. The basic separation element

in the cone concentrator is an inward sloping 117°) two

metre (6*25 feet) diameter cone. The pulp flow is not

restricted or influenced by side wall effects which occur

with the pinched sluice system. However, inched sluices,

also known as trays* are used within the cone concentrator

in certain applications with small tonnage products*

Feed pulp is evenly distributed around the periphery of

- 257 -

the cone. As the pulp flows towards the centre of the

cone the fine and heavy particles (concentrate) separate

to the bottom of the film* The concentrate is removed

by an annular slot in the bottom of the concentrating

cone; the part of the film flowing over the slot is the

tailings* The efficiency of this separation process is

relatively low and is repeated a number of times within

a single machine to achieve effective performance.

As the feed required for RDCC should be of 60 - 65%

solid with less fines making the feed close range

particles. It can be seen that nearly 35% of the total

solids are lost in this fines which contains 33% uranium

distribution. Recoveries obtained in RDCC are in the

range of 70 - 75% and further upgrading of this concentra-

te on wet concentrating tabling gave an average recovery

of 7Q%. The overall recovery obtained from this equip-

ment along with wet concentrating tables came out in the

range of 30 - 35%. Efforts were also made to recover

uranium values from the hydrocyclone overflow which were

accounting 33% of the uranium values by treating then

separately on wet concentrating tables. Our efforts

were in vain. There was hardly any recovery from this

slimy particles on the tables. This equipment was dis-

carded because it was giving an overall recovery of 35%

as compared to 45 - 50% recovery obtained from direct

tabling. Table 1 & figure 1 give a comprehensive idea

of the RDCC & its performance*

Ultimately, a decision was taken to discard the RDCC as

it was giving overall recovery, less than what was achi-

evable by tabling only. Then it was decided to further

expand the capacity of the Surda Uranium Recovery Plant

- 258 -

to match the capacity of the South Bank Treatment Plant

to treat tailings from 1000 MT/day copper ore treatment.

The recovery obtained from this plant ranges from 45 - 5556.

It is pertinent to mention here, that, initially the

Surda Uranium Recovery Plant was set up with 24 nos. of

wet concentrating tables to treat about 400 AIT of copper

tailings, i.e. 9 0.8 M.T of feed per hour per table.

It was later established that the recovery remained more

or less the same even if the feed rate was brought up

to 1*0 MT/hour/Table. Consequently, during the expansion

of this plant, only 16 nos* of tables were added to make

a total of 40 nos. of tables, to treat the entire 950 MT

of available tailings/( available from 1000 MT of copper

ore) per day.

DESCRIPTION OF THE THREE URANIUM RECOVERY PLANTS :

In Surda plant of UCIL which is treating 950 MTPO of

copper tails received from SBTP through gravitational

launders in a agitated tank from which it is pumped to

series of pulp distributors, thus distributing entire

tailings equally and giving 1 ton/hr. of feed per table

to 40 nos, of tables at a pulp density of 20 - 25%

solids. This provided 1 MT of feed/hr/table. The

concentrates obtained from the tables are collected

and pumped to the decantation pits where water decants

out and seal wet concentrates with, moisture of 10% is

transported to Jaduguda through trucks for further pro-

cessing for uranium recovery. Table tailings are

collected separately and pumped back to SBTP for sand

recovery for aines back filling. A layout of the

plant is given in figure 2.

Encouraged with the results at Surda Uranium Recovery

- 259 -

Plant, the Corporation took decision to set up a pilot

plant within the premises of Rakha Concentrator to test

feasibility of recovery of uranium concentrates from

copper tailings of Rakha concentrator plant utilising

4 nos. of wet concentrating tables. Results obtained

from pilot plant testing were quite encouraging and

gave an overall recovery of 40 - 45 % by feeding 0*8 MT

of tailings/table/hr at a pulp density of 20 - 25% solids.

Based on results obtained from pilot plant testing UCIL

set up another uranium recovery plant at Rakha with 48

nos of tables to treat entire copper tailings from Rakha

Concentrator plant of M/s HCL. Major modifications were

made in the layout of this plant to avoid the various

problems which were being faced in the Surd a Plant* A

layout of this plant is shown in Fig.3.

Detailed testings were also conducted at Mosaboni by

setting up a pilot plant with 4 nos of wet concentrating

tables initially which were subsequently increased to

8 nos to have more realistic testing trials* The concen-

trator plant of Mosaboni treats on an average of 2,700 MT

of copper ore/day contains uranium in the range of 0*007

to 0*008%* Testing results on this tailings on wet

concentrating tables were quite erratic and gave varying

recoveries due to which decision of setting up full scale

plant could not be made* The recoveries obtained were

quite erratic varying 10% to 30%. Reasons being, ore

from different mines (4 to 5 mines) are processed at

Mosaboni concentrator plant which also varies in basic

characteristics as well as in uranium values. Apart

from change in characteristics of ores the variation of

tonnages and mixing proportion of these ores caused

variation in the recoveries on which we were not having

any control* It was also observed that lot of uranium

- 260 -

Is lost In fines on tabling which could not be arrested

with the limited parameters available in wet shaking

tables.

Uranium distribution in the various size fractions of

the feed to the tables & of the table concentrates of

Rakha and Mosaboni and the comparative fractional

recovery of uranium at Rakha and Mosaboni are given in

table nos II & III.

It can be seen that the recovery from the coarser

fraction is more in case of MURP as compared to RURP.

This may be due to heavy gang minerals attached to the

coarser particles getting reported in the concentrate*

As evident from the table No III comparative fractional

recovery* that the recovery from fines is poorer at MURP

as compared to RURP, This is because uranium present in

the finer fraction are not recovered by tabling. It can

also be seen that recoveries from the liberted particle

sizes are less than 50* and minimum at MURP. At this

stage, it was decided to incorporate a " curved static

screen/Bartles Mozley Seperator/Cross Belt Concentrator

System (CTS/BMS/CBC system) at MURP because the copper

concentrator plant of Mosaboni, treating the maximum

tonnage of copper ore ie 2,700 MT/day, as compared to

1000 MT/day at the other two copper concentrating plants

at Surda & Rakha*

Initially 2 nos. of CTS & 1 no of BMS were installed in

the Mosaboni Uranium Recovery Pilot Plant. It was obser-

ved that 2 nos. of CTS were unable to give full feed to

the BMS* Also the BMS, even after being used in clea-

ning circuit, was unable to upgrade the concentrate to

- 261 -

the desired level. Thirdly, the feed taken for testing

of this system, was tapped from the main tailings disposal

line of M/s HCL/ICC & this gave erratic results due to

segregation of the tailings particles at the tapping

point*

At this stage the corporation, took decision to set up

a Tabling Plant with 32 nos. of tables to treat one third

of the total available tailings, i.e. 900 MT/day to start

recovering some uranium within a short span of about 2

years, pending a decision on the final process to be ~

followed for recovering uranium rrom the entire available

tailings. i*e* by tabling or by Direct Act leaching, or

by developing another alternative method in physical bene-

ficiation* This plant was commissioned in January 1987*

Next a decision was taken to install another 16 nos. of

tables & a balanced CTS/BMS/CBC system (with 3 nos* of

CTS, 1 no of BMS & 1 no* of CBC) for testing this system

under actual plant conditions* This decision materialised

by January 1988. A report on this CTS/BMS/CBC system

is given later on in this paper*

The Mosaboni Uranium Recovery Plant also adopted centra-

Used pumping system a9 adopted in the Rakha Uranium

Recovery plant. This minimised the pumping stages,

number of pumps & the power consumption. Besides, main-

tenance & inventory of spare-parts were also minimised

(by using similar pumps at the different stages of pumping)

A layout of this plant is given in Fig.4

Testing on CTS/BMS/CBC svst^" s

3 nos of CTS with 100 screens, 1 no of BMS, and 1 no of

CBC were installed in the main plant building of the

- 262 -

Mosaboni Uranium Recovery Plant, along with the ancilliary

equipment, viz agitated tanks, pumps, constant-head tanks,

compressor, pipeline etc. Testing on these equipment

started in June - 1988.

Two flowsheets were adopted as detailed in figures 5 & 6.

1. 1st Flowsheet J- Copper tailings equivalent to

the feed for 4 nos of tables were first taken to

the sump pit and through a sump pump, this was

pumped to 3 nos of CT5 evenly through a distribut-

or. The coarser fraction of the CTS was fed

to 2 nos of tables through a distributor*

On the tables, the concentrate was collected

and the table tailings was discarded* The

finer fraction of the CTS was fed to the BMS

via a surge tank* pump and a constant head

(giving 1*5 M head) feed tank* as recommended

by the supplier* Flush water to the BMS was

also provided through a constant head water

tank, placed about 2.5 M above the BMS feed

point, again as recommended by the supplier*

The BMS tailings was discarded totally through

a pipeline and gravity flow*

The BMS concentrate was collected in another

SRL agitated tank by gravity flow* This slurry

was the feed to the CBC and was fed by a pump

and a constant head feed tank* Water connec-

tions were given to the spray water pipes as

provided for by the supplier of the equipment*

The table tailings* The concentrate and the

fine middling were collected in a PVC tank at

"CBC Concentrate" whereas the coarser middlings

- 263 -

was recirculated back to the CBC feed tank by gravity flow.

All the three equipment, viz. CTS,BMS and CBC had been

installed on a platfoxm 5 M above the floor level. Whereas

all the tanks and pumps and also the tables were installed

at the floor level. This was to minimise the pumping

stages and to make maximum use of gravitational flow*

2. 2nd Flowsheet :- Here all the equipment were

kept in their same places as in the previous

flowsheet. Only the copper tailings was first

fed to 4 nos of tables through a distributor.

The table concentrate was collected. The :able

tailings was then pumped to the 3 nos of CTS

through the sump pump. CTS fines was as the

BMS feed, whereas the CTS coarse was discarded

with the tables tailings. The subsequent oper-

ation remained same as before*

The flowsheets as shown in figure* 7 & 8 show the material

balances also. Adopting flowsheet I. an overall recovery

of 30-35* was obtained, while adopting flowsheet 2, an

overall recovery of 35-45% was obtained. These are detai-

led in Table numbers from IV to XI.

Adopting flowsheet 1

Following were the observations in plant operation

In the CTS, about 15-18# of the fines was reporting with

the coarse fraction mainly due to fibrous foreign materials

in the feed, which reduced the effective screen surface

& also due caugulation of ultrafines with coarser particles*

in the BMS a feea pulp-density about \2% resulted in

- 264 -

Jamming of the decks and a density below 8% resulted in

improper bed formation on the decks. Because of excessive

heavies in the CTS fines (i.e. BMS feed) bed jamming was

a frequent phenomenon even at high slopes of 2.5° & high

orbital speed of 300 r.p.m.

Flowability on the Wet Concentrating Tables was poor &

a lot of wash water was required to prevent jamming of

the table decks* Increase in table slope & stroke length

helped little.

CBC deck was getting Jammed above 2056 feed density.

However lower densities remarkably improved the perfor-

mance of the CBC.

Adopting flowsheet 2

Following were the observations in plant operation s

Performance of the Wet Concentrating tables was normal*

CTS performance was better, since the big sized particles

(2-5 mm range) & the fibrous and foreign material in the

feed to our plant were removed on the tables. Deck

jamming was not encountered on the BMS decks since the

heavies were also recovered on the tables* As a result,

the BMS performed better. The observation on the CBC

was more or less a* in the previous case. However* bed

formation & separation were better.

FUTURE PROGRAMMES.FOR THE MOSABONI URANIUM RECOVERY PLANT $

The project to expand the existing tabling capacity of

the plant has been taken up* 48 nos more of tables are

to be installed to a make a total number of 96 tables,

which will handle the entire available Copper Tailings

- 265 -

from the Mosaboni Copper Concentrator, i.e. about 2700 MT

per day. Work on this has already been started & is expec-

ted to be completed by January - 1991. A salient feature

of this expansion programme is that a Thickener will be

incorporated in the circuit to dewater the table tailings.

As a result, the requirement for raw-water will be reduced

& hence the existing pumping capacity of raw-water from

the river to the plant (distance is about 2.75 Kms) is

not to be enhanced.

Apart from this, possibilities of recovering uranium by

Chemical Methods, is being looked into. The data collected

on extraction of uranium from the copper tailings by

chemical treatment route i.e. by the conventional acid-

leach process has conclusively shown that the recovery

by this method would be more than twice that by the

physical beneficiation route* A* on today, chemical

treatment it the only choice for maximum recovery of

uranium from the copper tailings.

The Control, Research & Development laboratory of M/s

Uranium Corporation of India Ltd., has carried out exten-

sive tests recently on copper tailings from the Surda,

Rakha & Mosaboni copper concentrators. These tests

have indicated that even from the Mosaboni copper tailings,

about 84# of uranium can be leached out with ferric

sulphate by the "low acid leaching technique1* developed

by them. Results of their study are shown in Appendix. 1.

It has already been established that with the copper

tailings from Surda & Rakha, leaching efficiencies would

be of the Sam* order. Thus one can safely assume an

average leaching efficiency of 80-89K for the copper

taJ.J.J.ng« from a}), the three sources.

- 266 -

For obtaining optimum leaching efficiency, the tests

were carried out under certain standard conditions. It

is known that during leaching, conditions for oxidising

uranium have to be maintained for its quick dissolution.

Normally, this is achieved by addition of commercially

available pyrolusite.

The same can also be achieved by addition of other

oxidising agents like sodium chlorate, or using ferric

sulphate solution itself as a leachant. Addition of

pyrolusite introduces the undesirable manganese ions

into the uranium leach liquor* Presence of manganese

in the solution make the final waste disposal procedure

more stringent. The tailings have to be neutralised to

pH of about 10.00 to complete precipitation of managenese

and ensure that does not leak out to the drainage

system* By using alternate oxidants this problem can

be solved* Use of ferric sulphate for leaching would be

ideal approach and this has an additional advantage*

The barren leach liquor can be recycled to the leaching

tanks after re-oxidation of the reduced ferrous ion by

bacteria*

The bacterial oxidation of ferrous iron has been taken up

in the laboratories at AMD and BABC and the concept of

recycling ferric solution for leaching uranium is being

looked into* However, experience to engineer this concept

into a plant scale operation'is yet to be achieved*

The major constraint for adopting the chemical process

for the recovery of uranium from copper tailings \» the

disposal of the leach tailings. Currently the copper

tailings are disposed off by M/s HCL on the banks of

Subarnarekha river close to the concentrator sites*

- 267 -

Environment agencies have exerted pressure on M/s HCL to

stop this practive. M/s HCL is planning to build a

tailings disposal system since last several years. So

far they have not been able to acquire 100 hectares land

for tailings disposal a small distance away from their

South Bank Treatment Plant. It will still take more

than 5 to 6 years for them to build and commission

tailings disposal system.

To incorporate chemical process and finalisation of

project report the following studies have been taken up

by us :

1. Finalisation of process parameters for bacterial

oxidation of ferrous to ferric*

2. Studies for loading characteristic of low

value of pregnant liquor and employing Elux

process in place of convention of ion exchange

system. This is being done to avoid chloride

iron build up*

3. Studies on environmental impact of the process

and tailings disposal*

In case our studies shows that chemical process will not

make much effect on the environment i.e. seepage of redium

remains below permissible limit and avoiding pyrolusite

oxidant, the major constraint of tailings disposal will

not be a problem in taking the decision for adopting

chemical process for recovering uranium from copper

tailings. As Indicated earlier that this process*.will

give more thanv. double uranium concentrate as compared

to physical beneficiation process and by adopting

- ?68 -

chemical process, the contribution of uranium mineralfrom copper tailings will be quite significant for nationalrequirement*

EXPERIENCES IN THE URANIUM RECOVERY PLANTS & THE HIGHLIGHTSTHEREIN X

Use of High Density Polyethylene Pipes :

For the first time in this company, " high-densitypolyethylene pipes'were used in the slurry lines in placeof the conventional rubber-lined pipes at the SurdaUranium Recovery Plant. Not only the cost of the H.D.P.Epipas wara 19*9, but they were also easier to installeasier to maintain & have a very long life* Initiallythese pipes were installed with rigid supports & clamps.But these pipes have a co-efficient of linear expansion,seven to eight times more than that of steel* As aresult, with tha fluctuation in temperatures, the pipaswara getting cracked* These pipas wara then laid looselyon mild-steel trays wifth loose clamping. This gave thenecessary room for expansion of the pipas caused by thatemperature variations. This gave a Yery good result& the system is working virtually trouble-free evarsinca.This piping system has than bean adapted at tha Rakha &Mosaboni Uraniua Recovery Plants also.

RAW WATER SUPPLY SYSTEM *

Generally, Intake Wells with vertical submersible pumpsara installed at river-banks to pump water to tha plants*But in tha Surda Plant, a sliding platform with a Centri-fugal Pump mounted on it, was installed at tha river-bankto pump water to tha plant. With tha laval of tha river

- 269 -

rising or falling, the platform, mounted on rails, could

be moved up or down with help of a winch. This system

was novel, extremely economical, simple, and efficient*

This system of installing sliding platforms at river

banks, in place of Intake Wells, has since been adapted

at the Rakha & Mosaboni Uranium Recovery Plants also.

MODERNISATION OF SUflDA URANIUM RECOVERY PLANT ;

After facing several problems in running the plant &

maintaining it, at Surda (which had virtually 2 nos of

tailings pumps & 2 nos of concentrate pumps for a

batch of 8 nos of Tables, apart from the other pumps)

the concept of Centralised Pumping System was brought

about in the Rakha Uranium Rec very Plants. These

Tables were installed in 3 bays at 2 different levels.

All the concentrate & tailings from the 48 nos of tables

were collected in drains with proper slopes to void

settling in them, and channalised to centralised pits,

froa where single-stag* pumping was used to pump out

both the concentrate and the Tailings. By this, the

number of pumps were reduced drastically. Maintenance,

down-time, & spare-parts costs of the pumps & their

inventory were minimised. Since the number of pumps

were less, the power consumption was also brought down.

This system has since then been incorporated in the

Mosaboni Uranium Recovery Plant also. Layouts of AURP

& MURP are given in figures 3 & 4.

By the time a total of 48 nos of tables were decided to

be installed in Mosaboni Uranium Recovery Plant, it was

felt that the Surda Uranium Recovery Plant had serious

flaws in its plant & equipment design & there was a lot

of room for making design & equipment alterations in the

plant. A decision to this effect was taken and a Moder-

- 270 -

nisation plan for the Surda plant was taken up in

September 1987 at a cost of about b.12 lakhs. The wet

Concentrating Tables were left untouched. Slurry fee-

ding system, tailings collection & disposal & concentrate

collection & pumping systems were changed & made simpler.

The number of pumps were brought down from 32 to 15, Out

of these 15 pumps, only 8 pumps are operated to run the

plant, rest are standby pumps. Here again the concept

of Centralised Pumping System was utilized. Power

consumption was brought down by about 3356. Since there

were fewer number of pinps, maintenance down time, man-

power required for maintenance, consumption of spare-

parts were all brought down. In fact, the saving on

power it-self offsets the cost of modernisation in about

3 years. The revised layout of the Surda plant after

modernisation in shown in Fig.9.

Our efforts are still on to recover the valuable

uranium from the wast* streams of Copper tailings.

This goes a long* way to minimise radiation & other

pollution hazards from these waste streams apart from

giving the country a vital atomic mineral, necessary

to implement its nuclear programmes.

•.ooOoo*.

Ml.?

\"7

LAYOUT OF THE SURDA URANIUM RECOVERYPLANT BEFORE MODERNISAT ION

I 6H*VPUIP P/STfilBurOR4 * 3 r ° K2 4:

J. b_S}4- 5 T/jfi'/NQS S PJTf

11- - •—nw

1

H

-1

1

H

1•1

^

—

-<

•i

h

1

7A7A

r

•

A7

K

1-

l rA

-e-

e

\

B

/

B

\

/B

\CO ID

/

B

\

/

B

\

B

/B

\

—B

\

/

\

\

Laa

\

/

\

/

\

/

V4Y4

4•

4Y

A

7A7A

I 0

7A7A7A

-\/\

/\aQ

/

\

/

\

/

\

\

/

\

/

\

/

J\f

\1

\I

a—

B

Q

B

(TJtea

9

B

&

R

-IB -

B

a

fi

- B -

4T4V4

4"V

4

-<

^

-t

-1

•«

-I

—

^

-41

1

CC*- =)

of

LA

YO

UT

A

TA

BL

E

h

1-

1

- 275 -

SFLOW SHEET FOR CTS/BMS/CBCSYSTEM TFSTING CIRCUIT NO- 1

rarrtK TAILINGS

IC.TS. SCREEN!

MO. » i rICT.SCOAHSEI

I CT.S-

COMC-

1 M SEPARATOR!

IIIOUS ER CONCl

__HJPPUf«D

IflWAL COHC. I

/*«

FLOW SHEET FOR C.TS/BM&/CBCSYSTEM TESTING CIRCUIT NO- *COfPfR TAII INGS

I TA B 'LE I " » I E TA t 5 i

I TABLE COWCl ICTS SCREEwi COARSCREJECT

I C T i f INES |

Ml '

I t M SEPARATOR^

I(ROUGHER CONO

DLINO 1

1

1 — I H TAILS REJECT

I { Q M C

I HHAL CONCI

MATEBIAL BALAWCt•'"I

t\. Ow SHEET — i

lOtPOO'lWOI

CIS. KRCtNl»O t*B I

icT.i cotim

| TttLE >—-

CT» FINES

COMClHTHtTt

t u n >—|K»3<IO»|ttU|

•

mm • » M»I««TO«|

•OUCMCR COHCll-IOIO^li I I I '97

1 C .

mni>t

• ' I t u,o, ••iijr r>-• I

»»coxet1 0-941

N

1r « 4 i fJl -H

MATERIAL BALANCE fQ» ' l O » SHEET - I

I'M.* I

En|4144>0«>*l«

™^Jr' I

|HI4|OBH'0 I

|H0|0-CHp4

I ' m i l . CONC1004

- 276 -

TAB IE -'w'

oO

1 .

2 .

3 .

4 .

5 .

6 .

7 .

s.

9 .

10

t 8 g «

Feed 1

Hydro

flow

: "Metal

to Hydro cyclone

cyclone over -

Hydroclone under f low

IIDCC <

RDCC (

RDCC '

8 Wet

Cone.

8 Wot

Tails

4 Wet

Cone.

i .4 Wet

Tails

:onc. 01

:onc. 0?

'a i ls

Tables

TC 1

Tables

IT 1

Tables

TC 2

Tables

TT 2

lurcical

Tonnesdry ore

1000.0

345.0

655.0

166.7

83.3

405.0

8 . 2

158.5

4 . 1

79.2

data on RDCC and Tables

Grade %U3O8

0.0104

0.0097

0.0107

0.0197

0.0169

0.0062

0.3055

0.0052

0.2347

0.0052

% Distributionof U3O8 at thestnee

100.0

32.7

67.3

31.2

13.5

?4.0

24.1

8»0

9 . 3

4 . 0

Kg U3O8at the •staee

104.00

33.90

70.10

32.SO

14.10

23.20

25.05

8.20

9.62

4.10

The overall recovery of uranium obtained was 33 % against 50 ?«

obtained with wet tables only. The major loss of about 33 % of

uranium was in the slimes, which went with Hydrocyclone overflow.

ooo 0 ooo

- 277 -

TABU _ 11

frpnfeed

ilAWAl

iiae

+4*

-18 -»-65

.- 65+1C0

-IOCK-O

-iscrroo-20Z+2.5

-?:S

.-'ceo ri-rtie :

I n -

1

3

?

1."

14

ro3S

'•:: «-.trli)utithe t,~!>le

ivt.

. 0

. 0

. 5

. 5

.C

. 5

. 5

..•

*"- •.

1 .

4 .

1 3 .

2 7 .

I-1.

o l .

100.

t .

0

0

5

C

0

•5

0

in

f rrc

0 .

0 .

0 .

0 .

0 .

w •

c.

•3.

xhe various size trrcxthe c -ncontr^te ,-rp n

In

CC1S

003?

0029

"03?

00" ?

0051

01,7

009 6

, U308

i . i

3 .1

5 .0

5 . ?

11.0

T... 4

C -> N•. Wt.

-

1.5

1 . 5

?.5

6 .0

10.5

30.5

•i7.5

Cane

ions of the!von bojiw,

c r. N "r. ..t.

,

1 . 5

3 . 0

5 . 5

11.5

•'.2.0

C-2.5

1CC.C

a A T £% U?O8

infrrcti-m

0.01T5

C.0286

0.0210

0.02S8

C.':292

C.030'.'

0.15«9

0.074C

"& U3O8

Uistributton

0.C5

0.57

0.7P

?.3O

4.OS

1.-.17

79.93

" ••p.l.;c5 in in* i Lider frrcti n» of« higher; tho vBluas in *.h« -3?5

ons ,-ro a l m s t rio ^lo *• i \*\a rwes of *.he food or Vhe concent>rtt.

T 5 0

••")

*1CC

t-150

T'JC

-3.-5

. '-t

l . C

' . 0

:'-\5

15.0

17.5

n.o

,. ..:• -..t

1 .0

•i.o

l ' .O

:'4.5

?i.5

57.0

100.0

infr?ott^n

0.0061

0.0-38

0.0047

O.OOSo

0.0050

0.0090

' . ' .COS?

'". L'.'-O?

- intr i -'-•itti'.n

l.'l?

1.95

S.51

9.46

12.10

14.10

55.40

15

•7

11

17

17

15

PO

V.t

. 5

. 0

. 0

.5

. 0

.5

• 5

Tone.

'• ',it

IP.5

IS. 5

T9.5

47.0

(54.0

71.5

100.0

Or ado

S A T :•.'.. V3JP

infvrctl-v.

0.C161

0.0.314

0.0?18

0.0??»0

0.0385

0.0673

0.10P0

0.06.?2

'.. U C8L>i»tri.' t:ti-"i

4.011.515.02

e.ie10.5?

16.77

53.30

- 278 -

TABLE - III

Comparative fractional recovery of Uranium atR.U.R.P and M.U.R.P.

Fractional recovery in %

Size R U R P M U R P

+50

+70

+100

+150

+200

+325

-325

52.0822.34

9.50

20.00

33.18

43.76

46.76

85.1123.16

30.69

25.91

26.18

35.77

28.96

- 279 -

TABUE

t g g/toE TESTPC (BIS

3 . 0 F.td 0 Co«™ 0 F inn B»S « « t . t ft DM cone. 0 DHS T*11.4 B>St»ce. C 1H hwttnHo. 6% UXB t * U30» A f««d i fine* C X U3OB 6 * O3Ct 0 * 4 Cytlt I Cop* i

• { t «U3M j I | { tM— t ft

1 . 0.0009 O.OOS O.C07* 50 0.0766 0.005 « . I J * M U . 2.9* 300

2. 0.0074 0.00M 0.0090 *0 0.0231 C.O06* 49.lt 5 ItaU. 2.5» 3000.0044 «

3. 0.00*4 O.0OS 0.01C6* W 0.O23T 0.0066 47.13 3 MlU. 2.S* . 300

0.01

4. C.0083 0.0047 0.009* ' 30 0.Q66 O.OOM S l .« 4 lUt*. 3.0* 320

5. 0.0091 0.0064 0.0111* SO O.0>73 0.0073 *60.2fl 4 MK*. 3 .0* 300

0.0103Q

6. C.0C73 e.ccm c.ocei* so c.ci2o o.co6i es.w 4 »»it«. a.s 3 0 0c.0094

Th> at. ff-*etlon of f i lm ind eoan* w»» fo-jnd to to SOX «rriox.

TABU N.. .

1ST CTAg TE T1W0 IKftWAiB. <t IKUt)

a.lto. F*t* to t*b|t* Cenetntnto Tall* % <O0t n*eo««» - Xcram f nctlon % UXt

I .

2 .

3.

4.

s.

&.

0.00a

0.0OS*.

0.006

0.0067

0.0064

O.OOS

0.04

0.CB04

0.044

o.o«a

0.03

0.030

0.004*

O.OOSO

0.0C49

0.0O7

o.oas

0.0C4

30.37

19.3

22.49

17.0

16.6*

30.00

- 280 -

1ST STAff TESTING (CBC FTfFOHWICE)

T A B U No. r . y l

a . * . « coc *"*J 1"3 Cone,

1 .

3 .

3.

4.

c cone. I coc T.m.« o cX « * 5 *I

ccc rtcov.nr 0% 66 S l o p . 0 O r b U t l B*lt

«p»»d

0.(266 0.0487 O.OOeB B.eyeltd tofMd

0.C231 0.O04 0.0372

0.0237 o'.OA8 0.0059

0.026* 0.0B2 0.00B9

0.0273 ,0.0504 0.0058

0.C120 0.0180 C.006 -do-

86.5

B0.3

86.1

7 7 . »

08.9

75

300 1.86

1.0

300 1.86 , ,

300 I-8* •»

320 X.96 „

300 i.eo , ,

260 1.9 , ,

S.Hs.1 UXM

2ID STAGE r-ffirg. nms

Ttbl» erne. n . % »feone.ntntt

TWbU TtlUJ H ! 3 S 2 _ _

TASLi IIP. V I I

T«bl« itcaviiyC

I.

I .

3.

4.

5.

6.

e.cces*0.OCT5

coon*• C.OOBf .

C.0105

0.0098

0.007

C.CW6

ErtUtttd

0.CM7

0.07*

0.0869

0.0496

0.06

0.O5J7

4.C5

3.16

3.66

4.81

2.58

5.17

0.0066'0.0071

0.0068

0.0083

0.0074

0.0051

0,O>K>

23.11

25.es

32.06

24.36

'22.03

25,3*

- 281 -

TADU .'u

Sl.tHi .

2 .

3 .

4 .

5.

6.

W

0.0 .

0 .

0 .

0 .

0 .

0

IVsT0066'0071

0068

0083

0074

0051

008

CTS

0 .

0 .

0 .

0

0

0

coars»s

00<9

005*

0069

0054

0047

.0071

STAGS r

cr : f intX U338

0.C075

0.0081

0.0087

0.00970.0071

0.0067

0.00980.0074

3TltC ( PERFOflWiCE OF BM3)

7T •it.it fln»t(Tab. ta i l

basis)

67.08

51.97

55.09

55.08

50.0

53.52

5Hi conc.l* U3J8 1

0.0134

0.0120

0.0137

0.0145

0.0110

0.0151

2:iS t a l lS U338

0.006'0.0039

0.0068

C.0072

0.0083*0.0071

O.C044

C0066

1I

BMS Hccot* r33

34

38.96

37.36

55.82

44.11

;ycl«tt in*!

4

4

4

4

3.3

4

Slontl

2

1.9

1.9

2

1.29

1.5

300

300

300

320

300

300

• Eitiaattd

TABLE JO.'

T-CTHO P.'.A:>:£ OH CSC)

ST. --- • • «N o . ',i-'-'S cane

i . Cone. <-3S I t i l CiJ - i tar i -i. ••3?.:\ * U336

•*arifa>icr»• lop* ursiial a*lt

1 .

2 .

3 .

4 .

5 .

6 .

3.9134

0.OI2C

0.0137

0.5115

0.0110

0.0151

• Eatlaatrt

00

0

0

0

0

0

.32*

.0233

,0351

,0207

.0213

.023

.029 '

0

0

0

00

0

0

.0357

.0058

.0072

.0074*

.0C53

.0061

.0091

ilecycl** \o

#

m

•

•

•

74.0

69.7

77.36

73.6

43.99

61.73

2

2

2

. 2

1

1

. 9

. 5

130

300

300

300

320

320

1

1

1

t

1

1

. 3 4

.83

.93

.83

.39

.39

n/nt.

•/at.

•/mt.

n/nt.

a/at

•/nt

- 282 -

1ST Stao* Ttttlna (Ovrall

TABUE NO.' X

5 1 . rctdNo. X U308

KCC. Lonbincdrtcovary(en flno»

)

o l a H«c.(coara* XTbaala) (tag*

Overafin* 11 reeovrv «• %

At coax**• tag*

Total

1. 0.0069 42.13

2. 0.0074 49.86

3. . 0.0084 47.12

4. 0.0083 91.92

9. 0.0091 90.24

6. 0.0073 99.69

86.9

80.3

86.1

77.69

88.90

7S.0

36.44

40.03

40.97

40.31

44.66*

41.76

23.39

19.30

22.49

17.0

16.68

20.0

20.06

24.34

23.83

24.04

=-.99

23.98

10.908

9.99

8.28

6.86

9.66

8.90

30.97

30.33

32.16

30.90

34.81

32.48

2io n x r Trarc levr MI tOFO v*wa)

TABIE Wo.

Owtmll wcawtv - 1lab K

S. tiH tt^r»U CDCr*c. CrUnrtric. TaMt»«.Ik. JHO08 * « (•» «!»• fc*tl») K M l i M labl*

•tap itw.

2S.it 13.69 25.11 ».O1.

3.

4 .

9 .

O.CCCS*0.007}

0.C0M*• . C O *

0.C1OJ

o.eow

0.CC7

33

34

- . 9 6

37.36

55.19

» . 7

77.36

70.6

tr.fi

3C.14

M.ca C.«

X.«3

M.0» I3.» 22.06 K.49

34.36 13.31 34.M 37.97

».O3 22,» 22.0S 44.U

APPENDIX 1

The fo l l owing t e s t s were conducted on laboratory s ca l e onMosaboni copper t a i l s :

1 . Low Acid Leach Test

The t a i l i n g s from Mosabani concentrator (feed to MURP)were c o l l e c t e d from time to time in batches (about 10 Kgs)and sent to the laboratory f o r leaching t e s t s . The wetsamples were f i l t e r e d , dried and mixed thoroughly. 10 Kgsample ( i n two batches) was taken for each t e s t .

Pulp density : 60* solidspH i during 1st hr 2.3

after 7 hrs 2.1 - 2.2Emf 450-500mV.

a) Using ovroluslte for oxidation

Sample No.and date

1.(25.

2 . (30 .

3. (06.

4 . (20.

5.(29.

6. (06.

3 .

3 .

4 .

4 .

4 .

5 .

88)

88)

88)

88)

88)

88)

Temp:C(Ambient)

29

31

36

30

32

38

Acidconsumption

17

15

14

15

10

13

b) Usina Ferric Sulohate

. 1

. 7

. 3

.3

. 0

.7

Pyrolusiteconsumed

Kg/T

5.20

5.65

5.00

5.25

4.00

2.60

Lea, china

HeadAssay

0.0074

0.0102

0.0071

LeachResidue

%u3o80.00195

0.0024

0.0015

0.00996 0.0023

0.0082

0.0Q92

0.0027

0.0017

LeachingEfficiency

73.6

76.5

78.9

76.9

67.1(pH 2.4 for8 hrs)

61.5

Head Assay t 0.0092%Pulp Densi ty : 60% s o l i d sTemp. : 36°CEmf t -450 mVH2S04 consumed t 10.4 Kg/TFe2(SO4)3 consumed t 6.5 Kg/TLeach Residue t 0.0014% ULeaching jtliciency : 84.8%

- 284 -

SIGNIFICANCE OF PETROLOGY IN URANIUM ORE PROCESSING

WITH SPECIAL REFERENCE TO THE COPPER TAILINGS OF

SINGHBHUM SHEAR ZONE

NP.SUBRAHMANYAM, T.S.SUNILKUMAR, D.NARASIMHAN

AND N.K.RAO

ORE DRESSING SECTION, BHABHA ATOMIC RESEARCH CENTREHYDERABAD

Petrology of the ore plays a key role andInfluences the technology and economics of theprocessing of the ores. In the absence of high gradeuranium resources, low grade ores constitute asignificant uranium resource in India, andpreconcentration before leaching is necessary inrendering-these resources viable, and in minimizingenvironmental deterioration. Copper tailings ofSinghbhum Shear Zone are such lean resources and UCIL isexploiting these ores by setting up preconcentrationplants at Rakha, Surda and Mosaboni. Intensivepetrological work has been carried out on these ores.Nature and distribution of uranium minerals is studiedby microscopic examination of thin and polishedsections: and the mineralogical composition anddistribution of uranium values in various size fractionshave been determined by a combination of sieving, heavymedia separation, radiometric assay and microscopicexamination. In the light of the petrological data,various problems involved in the preconcentration andleaching of these lean ores, and different technicaloptions in their exploitation are discussed.

INTRODUCTION

In planning for uranium ore processing, a

knowledge of the mineralogy and textures of the ore,

- 285 -

their variation and an understanding of the physical and

chemical behaviour of minerals and mineral assemblages

is essential. As mineralogy and textures in turn are

dependent on the genesis of the ore, petrolofiical

knowledge is very necessary for the process

technologist. Petrological work on the uranium ores of

Singhbhum in general and uranium bearing copper tailings

in particular is incorporated in this paper, and the

implications of the data on the processing of the ores

is dealt with.

PROCESSING OF URANIUM

Uranium normally occurs in minerals from which it

can be taken into solution by chemical means with a high

degree of selectivity from its associated gangue, with

high recovery. Chemical hydrometallurgy is

predominantly resorted to process uranium ores due to

this important property. An ore which does not meet

this criterion will require preconcentration by physical

beneflclation as in the case of Radium Hill. Similarly,

in the case of a low to very low grade ore, where direct

leaching may be techno-economically infeasible or

difficult, preconcentration by physical beneflciation

may make its exploitation feasible - as in the case of

by-product uranium - e.g., Palabora (IAEA Bull., 1980).

Economic feasibility of direct extraction of

uranium from the ore as well as preconcentration by

physical benefication is mainly Influenced by the rock

type of the ore, particularly its mineralogy and

texture. These factors determine the degree of

comminution required for the liberation of uranium

minerals and potential method for separating them by

- 286 -

physical beneficiation from the gangue. Mineralogical

composition also determines the nature of the lixiviant

required and the potential level of reagent consumption.

Copper tailings of Singhbum shear zone are very

low grade uranium resources, with uranium being

recovered as a by-product from these tailings.

Preconcentration before leaching has to be properly

evaluated in rendering these resources viable. A brief

history of the geology of the Singhbhum Shear Zone in

which these copper deposits occur is given here for

proper understanding of the host rocks and the nature of

occurrence of uranium minerals in them.

GEOLOGICAL HISTORY OF THE SINGHBHUM SHEAR ZONE

Decades of intense petrological research on the

Singhbhum shear zone (Banerjee,1969; Ghosh et al,1970;

Sarcar,1980; Rao.1977; Rao and Rao, 1983 a,b,c) has

indicated a complex history of the uranium deposits of

the zone, formed as a result of a continuous and

overlapping geological processes over a long period of

time. The process began about 2900 million years ago

with the emplacement of Singhbhum granite, considered to

be the geochemical source of 0 containing 7ppm U and

more or less culminated with the formation of economic

deposits of U, Cu etc in parts of the zone about 1500

million years ago. Various geological processes like

syngenetlc deposition with sediments, raetamorphism,

volcanism, orogeny and syntectonic granitizlation with

resultant mobilization and deposition of uranium by both

hypogene and supergene processes were responsible for

the economic concentration of the ore elements. These

- 287 -

diverse processes have left their imprint on the nature

of occurrence of the different ore minerals, their

association and their particulate characterisitics,

which have a direct bearing on processing for the

recovery of valuable minerals.

NATURE OF OCCURRENCE OF URANIUM IN SINGHBHUM SHEAR ZONE

DEPOSITS

Uranium in the Singhbhum shear zone deposits

occur in several forms; however uranlnlte is the

principal uranium mineral. From the textural features

atleast three types of uraninite have been recognised

(Rao and Rao,1983a), which represent different stages of

mineralization. These are 1)Uraninite I, characterized

by pitting and rounded or subrounded shapes, ii)

Uraninite II, a zoned type, generally idioaorphic,

always partly dissolved, giving irise to concentrically

arranged solution pits; this type not uncommonly has

often a core of the first type, and ill) Uraninite III*

an irregular type commonly associated with sulphides,

characterized by anastomir.ing irregular fractures which

are occupied by galena. This type not infrequently has

cores of the first two types. Besides, uranium also

occurs in the form of i) sooty pitchblende, ii)

secondary uranium minerals and surface coatings, ill)

uraniferous iron oxides, iv) U-Ti oxide — altered

brannerite and v) refractory uranium minerals such as

davidite, allanlte, sphene, xenotlme etc. In this

category of minerals uranium occurs as diadochic

replacement.

- 288 -

PROCESSING OF URANIUM FROM THE COPPER TAILINGS

Ore Dressing Section has carried out intensive

petrological and beneficiation studies on the copper

tailings of Rakha, Surda and Mosaboni (Degaleesan et

al,1967; Singh et al,1981, 1983 and 1985; Jha et al,

1987 and 1988). The host rock of mineralization in the

copper ores is quartz-chlorite-biotite schist, with

quartz and micaceous minerals being the main gangue

minerals. Apatite, magnetite, tourmaline and sulphide

minerals occur as accessories with uraninite as a trace

mineral, Mineralogical composition of these ores is

given in Table I. Because of the complex metamorphic

and metasomatic history of the host rocks, the main

uranium mineral uraninite is intimateley associated with

gangue minerals. Uraninite I and II are comparatively

coarse grained relative to uraninite III. The latter

occurs as small grained aggregates in the cleavage of

micas. Refractory uranium minerals occur as very fine

inclusions in the micas with pleochroic haloes around

them.

Uraninite has good physical properties which

should normally make it easily amenable for physical

beneficiation. Its specific gravity is 9.4 in contrast

to 2.66 and 3 respectively of quartz and micaceous

gangue. This property aids greatly in its separation

from the gangue by gravity methods. But the main

problem faced by the Mineral Engineer in physical

benefIciation of these lean ores is the differential

comminution property of micas with reference to the main

gangue quartz, and the behaviour of' the ore mineral

uraninite during comminution. Quartz and uraninite are

hard (H:7 and 5.5 respectively) but brittle. Micaceous

- 289 -

minerals have low hardness (H:2-2.25) but are

characterized by highly perfect basal cleavage, yielding

very thin, tough, flexible and v elastic laminae which

make it very difficult to grind them. While the Work

Index <WI) of quartz is only 12.77, it is 134.5 for the

micas, which is an order of magnitude higher.

The grindability of uraninite I and II are lower

than quartz, but that of uraninite III can be expected

to be much lower because of xhe inherent minute

fractures in it. Further a high proportion of Uraninite

III grains are intimately associated with the mica

minerals, in the cleavage of which they occur as

irregular fine dispersed grains. As the breakage

characteristics of the micaceous minerals and the

coarsely liberated uraninite are so vastly different, it

results in differential comminution and manifests in

non-uniform concentration of mineral values in various

sieve fractions (Singh et al,1981). Micaceous minerals

concentrate in coarser size ranges, whereas liberated

uraninite and quartz are concentrated in the fines. Due

to the differential comminution, uranium values will

have a bimodal size distribution In the ground material.

It is either in the cbarser fractions, enriched in

micaceous minerals occurring as unliberated uraninite

grains or In the finer size fractions due to faster

grinding of uraninite. Further grinding of the

micaceous minerals to liberate uraninites results in

more fine grinding of the already liberated uraninite

grains. Due to unliberated nature, uranium values in

the coarser fractions cannot be physically beneflciated.

Because of the very small particle size of uraninite in

the finer fractions, surface forces Influence greatly

and reduce the effect of gravity In their separation.

- 290 -

The problem becomes more acute with the increasing

micaceous content of the ore. Optimization of grinding

is necessary with cost analysis to get the best results.

To assist the mineral engineer in the

optimization of grinding, petrology laboratory has

evolved a simple procedure to estimate the liberation of

U-values in various size fractions and also to determine

the mlneralogical composition, by a combination of

sieving, heavy media separation, radiometric assay and

microscopic examination. The feed samples were sieved

into convenient fractions and subjected to heavy media

separation using bromoforn (S.G. 2.87) and methylene

iodide (S.G. 3.31). Bromoform lights (BRL), methylene

iodide lights (MIL) and methylene iodide heavies (MIH)

were obtained by this separation. Representative samples

from all these fractions were microscopically examined

and radiometricall; assayed.

Bromofom lights (BRL) mainly contain quartz,

whereas methylen* iodle lights (MIL) contain micaceous

gangue and apatite. Both the above fractions also

contain unliberated uranlnite. Methylene iodide heavies

(MIH) contain magnetite, sulphides and uraninite.

Uranium values in the MIH fraction can be taken as

fairly liberated and amenable for gravity separation,

whereas unliberated values in the BRL and MIL fractions

are not amenable. Data obtained from this study is

given in Table II - IV. In the light of this

petrological data, different options available for

processing of these ores are examined.

- 291 -

I.GRAVITY SEPARATION.

1. Rakha Copper tailings (RURP): A perusal of

Table II shows that the Feed sample contains 71.5% of

quartz and about 24% of micas. Uranium values in the

-140+270* BRL fractions are nearly completely liberated

and so the values in the finer fractions also can be

expected to be liberated. Hence, 9.16% of U values in

-270WBRL fraction are attributable to very fine

particles of uraninite which are prevented from settling

due to surface forces.

In the MIL fractions, which mainly contain

micaceous minerals, 8.87% values are locked up in +140

and 5.54% values in -140+270* fractions.. Out of the

7.65% values in the -270* fraction, some are again

attributable to very fine liberated particles of

uraninite.

In the MIH fractions, 11.63% liberated values are

reporting in +140* fraction and 19.86% in -140+270*

fraction.

The data shows that about 15% uranium values are

unambiguously unliberated from the micas. 32% U values

report in comparatively coarse liberated fractions and

are amenable to physical benefication by conventional

tabling. 53.6% values are reporting in fine sizes

(-270*) and these are difficult to be separated by

tabling. Fine grained gravity concentrators such as BMS

or CBC have to be used for efficient recovery of these

values. The ore apppears to be overground with respect

to uranium as high percentage of values are reporting In

fine sizes.

2) Surda copper tailings (SURP): Table III shows that

the feed contains 63.6% quartz and 31.6% micas. The

+1003BRL fraction contains 4.9% unliberated values,

whereas -100+2708 fraction has very little uranium

values . Micas contain a minimum of 18% unliberated

values. Out of the 70% fully liberated values 30% are

in coarse sizes and 40% in finer sizes. The values in

the latter can be effectively recovered by BMS or CBC

only.

3)Mosabonl copper tailings (MURP): The data in

Table IV shows that this ore contains 55% quartz and 38%

micas. Mica content of the ore is much more here in

comparison to Rakha and Surda. The effect of higher mica

content is to be seen in the high unliberated values of

about 43% in them. Out of the 50% liberated values, 20%

are in finer grained sizes. Hence, very low recoveries

only can be expected by tabling.

II.MAGNETIC SEPARATION

Uraninite is paramagnetic with a mass magnetic

susceptibility of 5 X 10 C.G.S. units and separable by

magnetic separation. Magnetite is ferromagnetic and

micas are paramagnetic and these will also certainly

report in the magnetics, increasing the bulk of the

magnetic fraction.

In Rakha copper tailings, as only 24% micas arc-

in the feed, magnetic separation may reduce the bulk to

a great extent for direct leaching. But experiments

have shown that magnetic separation by the presently

available WHIMS is not effective in separating fine

- 293 -

grained uraninite. As almost 53.6% U values are in fine

sizes in Rakha, WHIMS may not be reallly effective in U

recovery. HGMS or Superconducting magnets may be

helpful in separating these fine particles.

In Surda, magnetic separation may help in

reducing the bulk, but as almost 40% of the liberated

values are in fines, WHIMS here also may not be of much

help. In Mosaboni, as about 40% of micas are present,

bulk reduction here may not be much by magnetic

separation.

III. DIRECT LEACHING

Excepting for a small amount of apatite, copper

tailings do not contain any acid consuming minerals.

Hence acid leachants can be used for direct leaching.

Further, as much of the unliberated uraninite is in the

cleavages of micas, the leachants can easily penetrate

and salvage the uranium present in them, increasing the

recovery to a great extent.

But, while considering direct leaching of bulk

copper tailings, it should be noted that sizable amount

of uranium values are locked up in refractory minerals

such as allanite, xenotime, sphene, tourmaline, monazite

etc., and also as micro-unliberated grains of uraninite

in magnetite, micas and quartz. These values cannot be

recovered easily even by direct leaching. In the

comparatively high grade ores of Jaduguda, uranium

values contributed by these minerals may be of low

percentage, but in low grade copper tailings they

constitute a high percentage. These values are as much

- 294 -

as 33% in Surda (Singh et al,1983). Hence, recovery of

high percentage of U-values should not be expected by

direct leaching also. So, cost effectiveness of direct

leaching of bulk copper tailings vis a vis leaching

preceded by physical benefication has to be fully

evaluated. Relative effect of environmental degradation

by both the processes is also to be evaluated.

CONCLUSIONS

Uranium deposits of Singhbhum Shear Zone are

formed a3 a result of continuous and overlapping

geological processes over a long period of time and have

left their imprint on the mineralogy and textures of the

ores. The main uranium mineral uraninite occurs in three

different types in the copper tailings, out of which the

third type is intimately associated with micas, and has

Inherent fractures In it. Differential comminution

property of the micas is creating problems In the

liberation of U- values from the micas and also causing

overproduction of uranium fines. Conventional gravity

separation by tabling, or magnetic separation by WHIMS

are not effective in recovering uranium from the fines.

BMS and CBC separators for gravity, and HGMS and

superconducting magnets for magnetic separation may have

to be used for better recovery. The two options for

processing the uranium ores, i.e. (i) direct leaching

and (ii) preconcentration followed by leaching, have to

be fully evaluated in terms of cost benefit and

environmental degradation from the petrological and'

experimental data.

- 295 -

Acknowledgements

The authors express their sincere thanks to

Shri.P.R.Roy, Director, Materials Group, Bhabha Atomic

Research Centre for his sustained interest in the work

and kind encouragement.

REFERENCES

Armstrong F.C., 1974, Uranium Resources of the Future'Porphyry ' Uranium Deposits., Formation of raniumDeposits, Proceed. Symp. IAEA., Vienna, pp 625-635.

Banerji A.K., 1969, A Reinterpretation of GeologicalHistory of the Singhbhum Shear Zone, Bihar, J. Geol.Soc. India., v 10, pp 49-55.

Degaleesan S.N., Karve V.M., Viswanathan K.V.,Vijayakumar K. and Majumdar K.K., 1967, Report onBeneficiation of Rakha Mines Copper Ore (for NMDC),BARC/ Met / 10.

Ghosh A.K. and Banerjee A.K., 1970, On the Nature ofPetrogenesis of Dhanjori Lava near Rakha Mines,Singhbhum, Bihar. J .Geol. Soc. India, v 11, pp 77-81.

IAEA., Vienna, 1980, Significance of Mineralogy in theDevelopment of Flowsheets for Processing Uranium Ores.,Tech. Reports Series No: 196, pp 1-267.

Jha R.S., NataraJan R., Bafna V.H., Rambabu Ch. and RaoN.K., 1987, Recovery of Uranium values from CopperTailings of Mosaboni • Upgradation of BMS Concentrate onCross Belt Concentrator. A report submitted to UCIL.

Jha R.S., NataraJan R., Sreenivas T., Sridhar U. and RaoN.K., 1988, Amenability of Uranium Ores of Singhbhum toWet High Intensity Magnetic Separation. BARC/ 1-947.

Rao N.K., 1977, Mineralogy, Petrology and Geochemistryof Uranium Prospects from Singhbhum' She&r Zone. Bihar.Ph.D.Thesis, Banares Hindu University. ,

Rao N.K. and Rao G.V.U., 1983a, Uranium Mineralizationin Singhbhum Shear Zone, Bihar. I - Ore Mineralogy andPetrography. J. Geol. Soc. India., v 24, pp 437-453. ;

- 296 -

Rao N.K. and Rao G.V.U., 1983b, Uranium Mineralizationin Singhbhum Shear Zone, Bihar. II - Occurrence of'Brannerite '., J. Geol. Soc. India., v 24, pp 489-501.

itao N.K. and Rao G.V.U., 1983c Uranium Mineralizationin Singhbhum Shear Zone, Bihar. • IV - Origin andGeological Time Frame., J. Geol. Soc. India., v 24, pp615-627.

Singh H., Padmanabhan N.P.H., Rao N.K., Sridhar U. andRao G.V.U., 1981. Differential Comminution and itsApplications in Processing of Low Grade Uranium Ores.,Proceed Inter. Symp.on Beneficiation and Agglomeration,Bhubhaneshwar.

Singh H., Natarajan R., Das K.K.. Jha R.S., Sridhar U.,Rao N.K. and Rao G.V.U., 1983, Process EngineeringAnalysis of Uranium Recovery from Copper Tailings by WetTabling, BARC/ 1-771.

Singh H., Jha R.S., Natarajan R., Das K.K., Sridhar U.,Manmadha Rao M. and Rao N.K., 1985, Development of aGravity Concentration Process for Improving Uraniumrecovery from Copper Tailings, BARC/I-853.

Sarkar S.N., 1980, Precambrian Stratigraphy andGeochronology of Peninsular India : A review, Indian J.Earth Scl., v 7, pp 12-26.

- 297 -

Table I. Mineralogical Composition of Typical Feed (Wt%)

Mineral %

QuartzChloriteApatiteTourmaline

Opa-f Magnetiteques\ Sulphides

Others

SURDA ORE

62.322.32.33.63.25.8

0.5

RAKHA ORE

69.29.92.66.48.53.3

0.1

MOSABONI ORE

51.839.21.70.8

\J '0.9

Table II

F E E D

BRL(Quartz)

MIL(Micas*Apatite)

u r u

Win

Petrological Data on Rakha Copper Tailings

WeightX

*U3°8Distn.X

Weight*

Distn.X

WeightX

3

Distn.X

WeightXDistn.X

+140(>104pm)

45.141

20.5

35.7-

8.891

8.87

0.6111.63

-140+270(>52pm)

32.971

25.9

22.672

0.5

8.0662

5.54

2.1719.86

-270(<52pm)

22.0220

53.6

13.1363

(2)*9.16(0.3)

7.1996

(62)7.65(4.9)

1.6736.79

Total

100.090

100.0

71.5

9.66

24.05

22.06

4.4568.28

Figures in parenthesisvalues

indicate probable unliberated

- 298 -

Tablelll.Petrologlcal Data on Surda Copper Tailings.

F E E D

BRL(Quartz)

Apatite)

MIH

Weight*

Distn.X

Weight*eU3°8

Distn.X

Weight*

"U3°8

Distn.X

WeightXDistn.X-

+ 100

34.367

18.1

35.924

4.9

7.68170

10.28

0.722.92

-100+270

44.8104

36.6

27.372

0.43

14.6567

7.7.

2.7828.46

-270

20.9276

45.3

10.3543

(2)*3.49(0.2)

9.28120

(67)8.74(4.9)

1.2733.06

Total

100.0127

100.0

63.62

8.82

31.61

26.73

4.7764.44

Figures invalues.

parenthesis indicate probable unliberated

- 299 -

Table IV. Petrological Data on Mosaboni Copper Tailings.

F E E D

BRL(Quartz)

urrnXL

(Micas*Apatite)

M T Un In

Weight*eU3°8Distn.*

Weight*eU3°8

Distn.*

Weight*eU3°8

J o

Distn.*

Weight*Distn.*

+ 1000147pm)

40.286

30.9

27.6623

5.69

11.5162

16.65

1.048.57

-100+270(>52pm)

41.2102

37.5

21.345

0.95

16.6596

14.24

3.2222.29

-270(<52pm)

18.6190

31.6

6.8527

(5)*1.65(0.3)

10.30129

(96)11.89(8.8)

1.4516.06

Total

100.0112

100.0

55.85

8.29

38.45

42.86

5.748.92

Figures Invalues.

parenthesis indicate probable unliberated

- 300 -

IMPROVED GRAVITY FLOWSHEET FOR THE RECOVERY OF

URANIUM VALUES FROM THE COPPER PLANT TAILINGS.

R.Natarajan, R.S.Jha, U.Sridhar and N.K.Rao.

Ore Dressing SectionBhabha Atomic Research Centre.

The existing practice of preconcentration ofuranium values by wet shaking tables offers limitedscope for improving their recover/ from coppqr planttailings, particularly at Mosabani Uranium RecoveryPlant (MURP). The overall recovery at MURP is only18-22%. Extensive studies on improving the recovery ofthese values using fine gravity machines have beencarried out in the Ore Dressing Section laboratory andan integrated gravity flowsheet arrived at. A pilotplant using full scale machines was set up at MURP withthe help of UCIL engineers to test the suggestedflowsheet and collect data for the scale up anddesign factors.

The feed was classified into fines and coarsesizes using C.T.S 268 screens and the fines containinghigher distribution of uranium values was processed onthe Bartles Mozley Separator (BMS) and the Cross BeltConcentrator (CBC), while the coarse fraction wastreated on conventional wet shaking tables, supported bymatching conditioners and pumps.

The findings of the laboratory studies could notbe directly scaled up at the pilot plant stage due todissimilarities in the area of B.M.S and variation infeed characterstics, thus necessitating certain changesin the operating parameters of B.M.S. and furtheroptimisation studies of the same for maximising therecovery of uranium values.

The pilot plant studies have shown that an overallrecovery of 35-40% is feasible. This does not Includethe additional recovery obtainable by recoveringultrafine uranium values by hydrocyclones.

- 301 -

1.INTRODUCTION.

The copper deposits of Slnghbhum area are

uraniferous. The talling3 of the copper concentrator

plants at Surda, Rakha and Mosaboni, operated by

Hindusthan Copper Limited. constitute a significant

resource for uranium in India. The uranium content of

these copper plant tailings vary in the range of 90-120.

80-110 and 65-95 ppm respectively. Taking into account

the possibility of mixing of ores from other areas a

minimum assured average tenor of 90, 90 and 70 ppm can

be taken and at the current level of throughput at these

plants a minimum content of 111 tonnes per year of

uranium values is estimated (Table I).

Presently Uranium Corporation of India Limited

(UCIL) operates three uranium recovery plants at Surda

(SURP - 1000 TPD) Rakha (RURP - 1000 TPD) and Mosaboni

(MURP - 1500 TPD) using gravity concentration by Wet

Shaking Tables. The recoveries obtained in the three

plants of SURP, RURP and MURP are of the order of 40, 40

and 20%.

2. DEVELOPMENT OF AN INTEGRATED GRAVITY BENEFICIATION

FLOWSHEET.

A detailed analysis of the plant data has shown

that a considerable part of the uranium values occur in

very fine sizes (-400#)"'2>. In SURP feed 35% of the

overall weight is finer than 37pm (400») and contains

60% of the uranium values. In RURP feed 23% of the

material containing 47% of U^Og values is finer than

37,ui» (400*). In MPP feed the enrichment in fines Is the

highest with 25% of the material containing 59% of the

values. Further, studies on variation of recovery with

size during1 Tabling has shown that optimum recovery in

- 302 -

size lies in the range of about 74 to 37pm (-200+4001*)

and on either side there is a sharp drop. The recovery

drop in the coarse size range is due to nonliberation of

uranium values, while that in the finer size range is

due to the limitation of the shaking tables to recover

particles in this size range. The applicability of

different gravity equipments for efficient separation in

the different size ranges is given in Table II. Any

gravity equiupment would work more efficiently if the

feed is preclassifled into appropriate close size range.

It is imperative therefore that to improve overall

recovery of uranium values it will be necessary to aim

at improving recovery from finer sizes by using

appropriate equipment after preclassifying the feed.

Extensive studies were undertaken in the Ore

Dressing Section on the application of fine gravity

machines like Bartles Mozley Separator (BMS) and Bartles

Cross Belt Concentrator (CBC). These equipments have

proven their applicability in the recovery of tin,

tungsten and Nb-Ta mineral values in the fines size

ranges in many plants. The laboratory studies have

culminated in the development of an Integrated gravity

beneficiation flowsheet (Fig. 1)(2'*\ The process

involves the classification of the feed slurry into

coarse and fine streams over a CTS wet stationary screen

fitted with 100pm nylon sieve cloth. The coarse stream

is processed on conventional wet shaking tables and the

fines streams on BMS. The BMS concentrate is further

upgraded on another BMS or on a CBC. Uranium values

occurring in ultrafine sizes (<5/jm) that are not

recovered in the Bartles machine efficiently due to

limitation of the equipment are recovered using a

hydrocyclone as overflow.

- 703 -

3. LABORATORY TESTS.

The above flow sheet was tested in the laboratory

on the Mosabonl copper tailings using a CTS 216

screenbox with varying aperture nylon screen clothe3 for

classification and a semiautomatic BM Separator with 4

fibreglass decks. The results of these investigations

are summarized in Table III. These investigations have

shown that considerable improvement in uranium recovery

values, of about 50% could be achieved and a concentrate

assaying +500ppm UgOg from a feed containing 90 to

lOOppm U,OgCould be obtained. This, however, required

strict control over the classification as well as the

operating parameters of BMS. The optimum operating

parameters determined were orbital speed 210-230 rpm,

feed flow rate 75-88 x 10 a»9/s/m2, slope of decks

1-1.5°, cycle time less than 10 minutes and pulp density

10-12% solids. A stage of cleaning of BM rougher

concentrate on BMS was also found necessary. The ultra

fine values were recovered in a 75mm dia cyclone as

overflow.

4. LARGE SCALE TEST WORK AT BGML.

Having established the feasibility of the basic

flow sheet and the range of operating parameters several

large scale tests were carried out at the site of

Balaghat scheelite recovery plant at Kolar where a full

scale BMS was available. The main aim of these

experiments was to study the separation behaviour of

copper tailings on a full scale BMS and utilise the data

collected for scale upM>-

Though the overall results obtained were poor,

mainly due to limitations and mismatching of equipment

used for claasification, the test results showed the

- 304 -

efficiency of BMS in recovering values from finer sizes.

The BMS separation stage resulted in a stage recovery of

45% at ER of 2.9 in about 17* weight collection. These

tests on fullscale BMS gave valuable data on the effect

of various parameters of BMS operation on recovery c_

uranium values, viz. deck slope, drain time, cycle time,

orbital speed (rpm) etc. These tests also demonstrated

satisfactory scale up of BMS performance from laboratory

unit to full scale machine, and the reproducibility of

results under given operating parameters.

5. ON SITE TESTS.

Based on these findings it was proposed to

carryout further test work at the site of Mosaboni pilot

plant using full scale machines and drawing fresh slurry

from the main tailings disposal line from the Mosaboni

copper concentrator plant ' . The principal scaleup

parameters considered for the test work are given in

Table IV. A continuously operated pilot plant facility

was set up at the site. The main equipment included

CTS-268 screenboxes and a full scale Mark II BMS

imported by AMD. The other matching equipment like

conditioners and pumps were provided by ODS and UCIL.

The main pipe of 250mm diameter carrying the HCL

copper tailings to the disposal cyclone was tapped with

a 100mm dia pipe and the slurry brought to the receiving

sump of Mosaboni pilot plant. This was pumped up to a

two way distributor which fed the two CTS-268 screens.

The fines and coarse streams were collected into two

separate conditioners. These streams were made up to

required percent of solids by addition of water and fed

on to BMS and wet shaking tables respectively. During

the initial phase of test work problems cropped up in

- 305 -

maintaining constant slurry flow rate and pulp density

to the classification circuit due to fluctuations at the

HCL end. This resulted In blinding and poor

classification in the CTS screens and enough feed to the

BMS was not obtained. These problems were overcome by

adding make up tanks and pumps during the second phase

of test work. The schematic layout of the modified test

set up is given in Fig. 2. The feed classification and

the distribution of uranium values in the CTS streams is

given in Table V and the results obtained with BMS in

Table VI. It was observed that higher feed flow rate,

higher slope of decks and higher orbital rpm led to low

recovery of values in the BMS concentrate but at

increased ER. It was tehrefore necessary to optimise

the various parameters to yield concentrates with

optimum grade and recovery. Further the pulp density of

the feed to BMS was found to play a crucial role in

determining the separation efficiency of BMS. Optimum

stage recoveries of 75% with enrichment of about 3-4

could be achieved if the X weight collected in the BMS

concentrate is about 20-25%. The tests have indicated

the optimum parameters to be: slope 2.5°, rpm around

225, feed slurry flow rate 400 lpm ±8%, pulp density

10-12% solids, cycle time 6 minutes, drainage time 10

seconds and wash cycle 25-30 seconds.

6. UPGRADATION OF ROUGHER BMS CONCENTRATE ON CBC.

During the onsite test programme few experiments

were carried out to further upgrade BMS rougher

concentrates on the BMS itself. But the constraint of

having only one BMS made it Impossible to have

continuous runs. Further the percent heavies in the

concentrate being greater than 5% flow characteristics

of feed slurry on the decks changed in the cleaner

- 306 -

stage. This called for a new set of B M parameters td

be studied. Tests carried out earlier in the

laboratory3> had shown that BMS preconcentrate can be

efficiently upgraded in a CBC with high stage recovery

(75-80%) and ER of 2-3. Hence the rougher BM

concentrates were transported to ODS laboratory and

detailed tests were carried out on their upgradation on

CBCcts>. The levels of CBC parameters studied are given

in Table VII and the results in Table VIII. The results

demonstrated the reproducibility of earlier laboratory

findings.

7. TABLING OF COARSE STREAM.

The tabling of coarse sand3 being well

established no detailed tests were carried out for

parameter optimization. In a few tests carried out at

site, a stage recovery of 25-30% with 400-600 ppm

U«0owas obtained which was comparable to earlier results

obtained in the laboratory'2*.

8.ULTRAFINES RECOVERY USING HYDROCYCLPNE.

Tests on recovering very fine uranium values using

hydrocyclone could not be carried out at the site due to

the nonavailability of high capacity pump3. But from

the experience of tests carried out in the laboratory

and at BGML'2'"*' it can be said with confidence that an

additional 3-4% of Uranium values can be recovered using

small diameter hydrocyclones,operated at a dso of 5-7pm.

9. FURTHER MODIFICATION.

Based on the experience and the results obtained

during the on site tests at Mosaboni, it was found

desirable to add one more CTS-268 screen and also a CBC

- 307 -

machine for the upgradation of BMS concentrate. A new

pilot plant incorporating these machines has been set up

by UCIL at the MURP 3ite. and further test3 were carried

out by UCIL engineers. Experimental results of some of

these tests are included in Table IX. The results show

the feasibility of recovering about 35% of uranium

values at a combined grade of 450-500 ppm, from a feed

of tenor 70-85 ppm UsO under optimum conditions of

operation.

10. DISCUSSION.

The overall results obtained in the series of

onsite tests, while giving substantially improved

recovery over direct tabling, fell short of results

obtained during laboratory tests. While the laboratory

tests predicted an overall recovery of about 45% at

about 500 ppm grade, from feed assaying 70-95 ppm

(excluding ultrafine recovery by hydrocyclone), the

onsite pilot plant scale tests proved the feasibility of

recovering 35% of yalues at 450 ppm grade with high

confidence level. A few tests, however, gave upto 40%

recovery, though at a somewhat lower ER. An exhaustive

analysis of the results obtained was carried out to

pinpoint the reasons for inferior results and the ways

of further Improvement. The findings are discussed

below:

10.1.Liberation analysis: Liberation of vulues was

evaluated by heavy media separation. The onsite test

(OST) feed samples w*re separated into different density

fractions by using Bromoform (S.G.2.81) and Methylone

Iodide (S.G.3.31) liquids. The U*O contents of the

fractions so obtained were assayed radlometrlcally and

the distribution of uranium values in each fraction

calculated (Table X). It is seen that the. OST 'samples

- 308 -

have higher distribution of uranium values in the quartz

gangue and micaceous minerals. This poor liberation can

be attributed to relatively co»rser grind now practised

at the Mosaboni copper concentrator plant.

10.2.Feed Assay: The UaOa assay of the feed samples

collected during different on site test work showed wide

variation from 65ppm to 95ppm over the period of test

work. On the whole the feed assay is lower as compared

to laboratory test samples.

10.3.Size distribution in the feed: The size analysis of

the OST feed samples has shown that about 15% of the

material is coarser than 65* and 6-10% is coarser than

48». The material finer than 2001* is only 35-42 as

compared to the lab sample which contained 45-50%

passing 200tt. It is apparent that OST samples are

relatively coarser, which is reflected in the lower

distribution of liberated uranium values in the MIH

. fraction and higher distribution in the BRL (quartz) and

MIL (micaceous) fractions.

10.4.Scale up Criterion: The laboratory investigations

for the development of the integrated gravity flowsheet

were carried out on a Bartles CTS 216 screen box, a

semiautomatic laboratory model of BM Separator and a

full scale Cross Belt Concentrator. The Industrial CTS

268 screen box Installed at the Mosaboni site* is wider

in the direction of slurry feed flow compared to the

Laboratory model. Based on laboratory test work a

throughput of 2.3 TPH/m was suggested by ODS for scale

up. In view of the higher mica content in the Mosaboni

tailings resulting in higher blinding rate. this

throughput was observed to be on the higher side.

Bartles engineers have suggested a throughput of about 2

TPH after evaluating feed characteristics.

- 309 -

The laboratory model of BMS has 4 decks each

1.22m long and 0.76m wide while the BMS full scale

machine has 40 nos of 1.53m long and 1.22m wide decks

each of which is divided by a central spacer along the

length of the deck resulting in a change of geometry of

decks compared to the laboratory model. Any change in

flow rate will change the velocity of the slurry on the

decks but not the film thickness. Due to geometric

dissimilarity the performance of the laboratory model

BMS can be duplicated on the full scale machine only by

changing the kinematic and dynamic conditions of feed

flow. Kinematic similarity can be achieved by changing

flow rate of feed on the decks and the deck slope. The

dynamic similarity can be maintained by varying feed

flow rate and the orbital rpm. The flaky nature of

micaceous minerals tended to occupy more surface area on

the decks inhibiting free flow of material, The higher

distribution of values In the mica fractions in the OST

samples compounded the problem.

10.5.BMS performance. Higher throughput to BMS beyond 3

TPH by increasing the pulp density has shown cake

formation on the decks resulting in poor performance.

An optimum weight collection of about 20% as concentrate

is necessary to yield maximum recoveries at optimum

grade. Efforts to improve the enrichment factor by

reducing the weight percent collected in the concentrate

led to low recovery.

10.6.Tabling of coarse stream: The shaking tables

presently employed at MURP have slime decks, which may

not be Ideal for processing the coarse stream. In view

of the changed characteristics of feed to the table, a

re-evaluation of the parameters of operations of the

shaking table, particularly flow rate, julp density.

- 310 -

quantity of wash water and slope of deck may be

necessary to achieve optimum results.

11.CONCLUSIONS.

1. About 35-40% of uranium values at an enrichment ratio

of 2.5-3 can be recovered in the BMS concentrate.

2. The BMS concentrate can be further upgraded on Cross

Belt Concentrator at an ER of 2-3 and a stage recovery

of 65-80% depending upon initial grade.

3. Feed characteristics such as grade, grind, percentage

of liberated uranium values and the mica content have

pronounced effect on the recovery values.

4. It should be possible to achieve a combined

concentrate with about 40% recovery at a grade of 500ppm

U3Og from the Mosaboni tailings of tenor 70-80 ppm by

adopting integrated gravity flowsheet under optimum

conditions of operation. With higher tenor of feed

higher recoveries should be feasible.

ACKNOWLEDGEMENTS.

The sustained support and encouragement extended

by Shri.M.K.Batra, Chairman and Managing Director, UCIL

during the course of the studies, both in the laboratory

and on site is gratefully acknowledged. The authors

also thank Shri.K.K.Berl, Shri.U.K.Tiwari and

Shri.J.P.N.Lai of UCIL for their cooperation and

involvement during the teat work. The authors are

grateful to Dr.M.V.Ramaniah, former Director,

Radiological Group for his sustained interest, and to

Shri.P.R.Roy, Director, Materials Group for his

continued encouragement in the investigations.

- 311 -

REFERENCES.

1.H.Singh, R.NataraJan, K.K.Das, R.S.Jha, U.Sridhar,N.K.Rao and G.V.U. Rao. "Process engineering analysisof Uranium Recovery from copper tailings by WetTabling". BARC IR 2-771 (1983).

2.H.Singh, R.S.Jha, R.NataraJan, U.Sridhar, M.M.Rao,K.K.Das, N. P. Subrahmanyam and G.V.U. Rao. "Laboratoryinvestigations on improving Uranium recovery from coppertailings by gravity beneficiation". InvestigationReport submitted to UCIL (1984).

3.H.Singh, V.H.Bafna, R.NataraJan, R.S.Jha, U.Sridhar,Ch.Rambabu and N.K.Rao "Uranium Recovery by improvedgravity concentration process incorporating BartlesMachines- Bartles Mozley Separator and Bartles CrossBelt Concentrator". Report submitted to UCIL (1985).

4.R.NataraJan, R.S.Jha, K.K.Das, U.Sridhar, H.Singh,Ch.Rambabu and N.K.Rao. "Large scale experiments at BGML(Kolar) on beneficiation of Uranium values from Mosabonicopper tailings". Investigation Report submitted toUCIL, March 1985.

5.R.S.Jha, R.NataraJan. V.H.Bafna, K.K.Das, U.Sridhar,N.P.Subrahmanyam, Ch.Rambabu and N.K.Rao "Recovery ofUranium values from copper tailings by Gravity Process:Onsite tests and scale up studies at Mosaboni Plant".Report submitted to UCIL. February 1987.

6.R.S.Jha, R.NataraJan, V.H.Bafna, Ch.Rambabu and N.K.Rao"Recovery of Uranium values from copper tailings ofMosaboni: Upgradation of BMS concentrate on Cross BeltConcentrator". Report submitted to UCIL. April 1987.

7.S.Chakravorty. J.P.N.Lai, U.K.Tewari and K.K.Beri."Recovery of Uranium mineral concentrate from fineparticles". Paper presented at the Seminar on RecentDevelopments in Mineral Engineering. April 1989Jamshedpur.

- 31? -

Table I. Uranium in Copper Tailings.

I.Tailings at presentThroughput TPD.Million Tonnes/year.

2.Tenor U_O_ppm0 0

RangeAverage

3.Contained U,Og in ions.

At min. grade TPYAt average grade TPY

4.Contained U-Ogln tonnes

Per million tonnes ofore processed.At minimum gradeAt average grade

Surda

0.3

90-120105

2732

90105

Rakha

iOOC0.3

90-110100

2730

90100

Mosaboni

27000.8

70-10085

5760

7085

Total

47001.4

111131

Table II. Distribution of U Values in different sizes inMosaboni tailings and and applicability of gravitybeneflciatlon equipment.

Mosaboni % weight

Wet shaking tableB.M.SeparatorCross Belt ConcentratorHydrocyclone

Particle size range (MDI)

>100

16

YesNoNoNo

100-37

26

YesYesYesNo

37-7

48

NoYesYesYes

<7

10

NoNoNoYes

Table III. Laboratory Test Results Using IntegratedGravity Flowsheet.

Lab. test code

Feed assay

BMSRGHRCone.

BMSCLNRCone.

DTCone.

MIXED• TABLECone.

CYCLONEOVER-FLOW

U.Cone.

Wt.%assayDist%

Wt.%assayDlst%

Wt.XassayDist%

Wt.%assayDlst%

Wt.XassayDlstX

Wt.%assayDistX

MSB-PP

100

17.634059.8

7.869053.8

2.7736710.2

10.5760564.0

2.51303.3

13.151567.3

MPP-1

95

19.022244.5

7.647037.6

0.636104.0

8.2348041.6

2.732106.03

11.041347.6

MPP-21

85

17.519039.3

5.3852833.4

2.06410

9.92

7.4449543.3

2.752006.49

10.241549.8

MPP-22

70

17.219040.7

5.041029.3

1.353010.2

6.343439.1

3.02008.6

9.336047.6

SI.No.

1

2

3

4

5 '

Table IV. Equipment

Equipment

CTS Screenbox capacityB.M.SeparatorcapacityPulp densityCross beltconcentratorShaking table

HydrocycloneSizePressure

CapacityType

Units

TPH/m

n»3/hr/m2

Xsollds

m3/hr/m2

TPH/nT

mmKpa

m3/h—

Scale-up

Lab Test

2.55

0.35

8.0

500 Kg

0.13

75.0100.0

4.0Dorrci

i

Criterion.

BGMLTESTS

-

0.34

11.0-

-

100.0150.0

10.0one

PLANTDESIGN

2.3

0.35

10.0

500 Kg

0.15

150.0340.0

20.0Rletma

Table V. Classification and Uranium Distributionin CT5 streams.

No.

1234567

Code- Nn

OST 2/3OST 4A/3OST 4B/3OST 6A/3OST 7A/3OST 8A/3OST 9B/3

Feed

TPH

3.53.33.34.15.15.04.0

U3°8

90858576907264

Fines

TPH

1.81.71.72.52.82.52.3

°3°8109100100961088977

Coarse

TPH

1.71.61.61.62.32.51.7

U3°8

70696945676054

Wt% inf 1 np<*

51.451.551.561.055.050.057.5

Dist.inf 1 r>A<t

62.360.660.677.066.058.369.2

Table VIB.M.S. Test Resultd Under Optimum Parameters

At Constant Deck Slope of 2.5° and Cycle Time of 6 Min.(All assays are by chemical analysis in ppm. )

CodeNo.

2/34A/34B/36A/37A/38A/39B/3

RPM

225225240225225225225

Flowrate(lpm)

428428420435420430460

Feed

TPH

1.41.91.92.52.272.32.02

U3°8

1091001009610B8477

Concentrate

TPH

0.340.370.430.340.480.520.37

WT%

24.019.422.413.421.122.618.0

U3°8

254254256335325203181

Recovery

Stage

55.249.357.047.063.054.242.3

Overall

34.530.334.536.141.631.629.3

Table VII. Levels of C.B.C. Variables.

Belt speed M/mln.Percent

Shear rate rpm.Feed slurry flow rate lpm

Upper

1.523024060

Lower

1.392522545

- 315 -

SI.No.

*12345676

Table VIII. Results of C.B.C. Tests(All assays are by chemical analysis in

Feed

U3°8

147159156150140140146140

X weight

24.122.539.414.033.821.351.052.2

Concentrate

U3°8

362414283545279420220212

X Dist.

59.458.671.850.867.463.977.179.0

ppm)

E.R.

2.462.601.823.632.003.001.511.51

Table IX. Results Obtained on Modified TestFacility at MURP (carried out by UCIL)

(All assays are in ppm)

Experiment No.

F

S

E

D

FINES

COARSE

BMS Cone.

CBC Cone.

Table Cone.

CombinedConcentrate

Wt.XassayDlstX

Wt.XassayDistX

Wt.XassayDistX

Wt.XassayDistX

Wt.XassayDlstX

Wt.%assayDlstX

1

50.07655.0

50.06245.0

6.026623.1

2.848720.0

2.040011.6

4.840031.5

2

50.09060.8

50.05839.2

11.923137.1

4.450429.8

0.885046.0

5.2850435.8

3

57.910068.9

42.16231.1

11.523732.4

5.13458X

28.0

1.54407.94

6.6345435.9

4

50.09959.6

50.06740.4

9.726630.9

3.2262224.1

1.254586.9

4.4757631.0

- 316 -

Table X.

LAB SIOST SIOST S2OST S3OST S4

Liberation

BFL

Wt.%

63.1761.6050.9062.0054.80

eU 3O 8

2047434537

%Dist

13.4630.8023.4033.2020.20

Analysis of Mosaboni Feed.

MIL

Wt.%

34. 1035.1043.7035.9041.00

eU3°8

9611710498117

%Dist

35.4143.7048.6041.9047.00

MIH

Wt.%

2.733.305.402.103.40

eU3°8

916727486996982

%Dist

51.1325.5028.0024.9032.80

Fig.l. Integrated Gravity Flowsheetfor Uranium Recovery from Copper Tailings.

Feed

|CTS

i

iWET

1SCREENS|

iFines

IBARTLES MOZLEY

SEPARATOR

RougherConcentrate

Coarse

1WET SHAKING

TABLE WILFLEYor DIESTER

Tailings DT Cone. DT Tails

BARTLESCROSS BELTCONCENTRATOR

-Tailings-

Middlings CleanerConcentrate HYDROCYCOLNE |-

Sl lines

-Sand-

MixedTable Cone.

1Final

Uranium Cone.

Final WasteTails

- 317 -

Fig. 2. M0SABAN1 PILOT PLANT SET UP

BUS FEEDBOX

Cu. TAILINGS

PULPOST

WILFLEYTABLE

CLASSIFICATION B.M.S. STAGE

TAILSPUMP

-REJECTS

COARSE TABLINGSAMPLINO POMTS

- 318 -

MAGNETIC SEPARATION FOR PRE-CONCENTRATION OF

URANIUM VALUES FROM COPPER PLANT TAILINGS.

R.S.Jha, T.Srsenivas, R.NataraJan,

U.Sridhar and N.K.Rao

Ore Dressing Section

Bhabha Atomic Research Centre.

Using the paramagnetic character of uraniumminerals, pre-concentration of copper plant tailings ofSinghbhum area have been investigated in a pilot plantscale wet high intensity magnetic separator (WHIMS).The experiments were aimed at maximising the recovery ofuranium values in the magnetic fraction, as it wouldgreatly reduce the quantity of the material to beprocessed by leaching and improve its grade to highereconomic levels.

The variables studied include magnetic fieldintensity, matrix drum speed, feed slurry flow rate andits pulp density. The results of these investigationshave shown that 75-85% of the contained uranium valuescould be recovered in 45-55% weight in the magneticfraction in the case of copper plant tailings fromRakha, Surda and Mosabani. The losses in thenon-magnetics were primarily due to the ultrafineliberated uraninite particles not collected by WHIMS dueto machine limitations and the values occurring as fineinclusions in quartz.

Improved recovery can be obtained by offeringhigher field gradients and preventing loss of very fineliberated uranium values. High gradient magneticseparator (HGMS) offers higher field gradients. A testin HGMS has indicated superior results in comparison toWHIMS.

. INTRODUCTION:

Copper plant tailings of Singhbhum (Bihar) has been

recognised as one of the significant resource of

by-product uranium in India. The average tenor of the

tailings from the three copper plants at Surda, Rakha

- 319 -

and Mosabani operated by Hindustan Copper Ltd. are in

the range of 100-130 ppm, 80-100 ppm and 70-100 ppm u»0«

respectivly. The lower tenor of these tailings impose

various techno-economic constraints for its direct

leaching for recovery of uranium values. Gravity

beneficiation plants using wet shaking tables are in

operation for pre-concentration of uranium values from

these tailings. However, the recovery of uranium values

in the concentrates of the gravity plants are only

20-25% in Mosabani and 35-40% in Surda and Rakha(1) due

to complex mineralogical composition and limitation of

shaking tables in recovery of values from finer size3

(finer than 53 t-tm). Integrated gravity concentration

flow-sheet developed in ODS can improve the recovery of(2)uranium values considerably . Other physical

beneficlation methods having potential for further

improving the recovery are flotaion and magnetic

separation. Magnetic separation in units like Wet High

Intensity Magnetic Separators (WHIMS) and High Gradient

Magnetic Separators (HGMS) have capabilities of

recovering extremeley fine particles of weakly(3 4)paramagnetic materials such as uranium minerals ' .

Results of studies carried out on recovery of uranium

values from copper plant tailings by magnetic separation

in WHIMS and HGMS are discussed in this paper.

2.PRINCIPLES OF WHIMS AND HGMS:

Magneitc separation techniques for separation of

minerals have been in use for many years. Recent

advances in magnet design have led to the development of

large WHIMS and HGMS. Separation of weakly magnetic

particles from diamagnetic or non-magnetic particles in

these separators is a physical separation based on three

way competition between magnetic forces , viscous forces

- 320 -

or drag forces and gravitational forces . The

magnetic forces pull the magnetically susceptible

particles in one direction for getting them captured on

the surface of the matrix material. The other two

forces pull all the other particles in another direction

and they also try to compete with the magnetic forces in

driving the magnetic particles in the same direction.

The magnetic force depends on the volume of

particle, difference in magnetic susceptibility of

particle and fluid, and on the magnetic field and its

gradient. The fluid medium normally used in idu3trial

practice is water and hence susceptibility difference of

partilcle and fluid does not remain an option to

increase the magnetic force on the particle to recover

it in the magnetic fraction. This is possible only by

increasing the applied field and its gradient which can

be produced In a variety of ways of magnet and matrix

design. In WHIMS the magnetic field is high and in HGMS

the gradient of field is also high due to design of

matrix. The magnetic force density in WHIMS and HGMS

are of the order of 4xlO9(N/m3) and 6xlO11(N/m9)

respectiveley. Superconducting HGMS can produce still

higher force density than conventional HGMS* .

3. EXPERIMENTAL PROCEDURE:

3.1.WHIMS'- Experiments were carried out in a continuously

operated pilot plant scale WHIMS. Random samples of

20-60 Kg drawn from the bulk sample received from UCIL

were pulped into slurry with 10 to 30% solids, and the

slurrry was fed into the WHIMS at a flowrate between

12-20 lpm in the different experiments. Separation was

carried out at magnetic flux density of 1.5 and 1.8

Tesla, achieved by regulating the current to the

solenoid of the electromagnet. Three fractions, a

- 321 -

magnetic (MAG). a middling (MID) and a noVi-magnetic

(NMAG) fraction were collected separateley. A few

experiments were also carried out in two stages of

•magnetic separation; first 3tage at 1 Tesla and NMAG of

first stage was scavenged at 1.8 Tesla.

3.2.HGMS: A few small scale tests were carried out at the

Sala Magnetics Division, Allis Chalmers Corporation, USA

with classified fines of Mosabani copper plant tailings

on a laboratory model HGMS. Four tests have been

carried out with varying matrices and flow rates, each

test in three sequential stages under applied magnetic

flux density of 0.5 Tesla, 1 Tesla and 1.5 Tesla.

4. RESULTS:

4.1.Copper Tailing3 from Mosabani Plant (MURP):

The results with MURP sample (Table I) showed

that the magnetic fraction assayed about 180 ppm U»O

having 87% ditfibution of uranium values in 53% weight

when the WHIMS was operated at 1.8 Tesla of magnetic

induction, matrix drum speed of 2 rpm and slurry density

of about 10% solids (code WH/3). Increase in pulp

density to 30% solids and drum speed to 4 rpm (Max.)

resulted In drop of grade of MAG mainly due to about 10%

higher collection of weight (code WH/2). Lowering

magnetic induction to 1.5 Teala showed a marginal

decline in grade and recovery (code WH/5 & 7) to 158 ppm

UsO and 80% recovery of values respectively with

weight collection of about 51%. In one of the

experiments (WH/9) the NMAG at 1 Tesla was scavenged at

1.8 Tesla and MAGS and MID of both the stages mixed

together gave a recovery of 94% in 70.2% weight.

The results of HGMS test with classified fines

(finer than 100 ^m) showed (Table II) that more than 90%

- 322 -

of the uranium values could be recovered in the magnetic

fraction in about 50% weight.

4.2.Copper Tailings from Surda Plant (SURP):

The results with SURP samples also showed similar

results as with MURP samples. A recovery of about

81-87% could be achieved in 50-56% weight. Scavenging

NMAG obtained at 1 Tesla at 1.8 Tesla reduced the value

in final rejects to 72 in 35% weight. The results are

shown in Table III.

4.3.Copper Tailings from Rakha Plant (RURP):

The results with RURP sample showed (Table IV) that

the recovery of values is limited to 75% in WHIMS even

at 1.8 Tesla of magnetic induction. At lower magnetic

induction of 1.5 Tesla and 1.3 Tesla the recovery drops

further by about 10% with almost same percent drop in

weight collection in magnetic fraction. Scavenging NMAG

obtained at 1 Tesla In WHIMS at 1.8 Tesla din not show

any significant increase in recovery of values but an

increase in weight collection by about 8%.

5. DISCUSSIONS:

5.1 Effect of Feed Characteristics:

Liberation studies of the test sanples of copper

plant tailings from MURP. SURP and RURP (Table V) showed

that about 42.8%, 26.7% and 22.1% values are composite

with micaceous minerals in the three samples

respectiveley in 38.4%, 31.6% and 24.1% weight. These

micaceous minerals are coarser In size compared to

uranium minerals due to their inherent flaky nature. On

the other hand they have lower specific gravity than

uranium minerals and hence they experience much lower

- 323 -

drag force and gravitational force compared to average

uranium particles. The magnetic suscpetibility of

micaceous minerals are also higher than the uranium(7)

minerals and hence they have much higher probabilityof being captured in the matrix of WHIMS and report in

magnetic fraction. This results in almost entire weight

percent of micaceous minerals reporting in the MAG

increasing its weight percent.

Quartz fraction of the three samples have only

8-10% values composite with it in 56%,64%, and 71%

weight of MURP, SURP and RURP samples respectiveley.

Though the quartz is basically diamagnetic, due its

partly composite nature with mica and other magnetic

minerals some amount of it is reporting in the MAG

fraction. Moreover, entrainment of coarse qurtz

particles in the matrix is also possible due to higher

loading of matrix resulting from high content of

magnetic materials present in the sample and thus

results in an addition to the weight percent of MAG

without corresponding addition in recovery of values.

The uranium values which are fully liberated or

composite with magnetite, a strongly magnetic mineral,

are 48.9%, 64.5% and 68.2% in MURP, SURP and RURP sample

respectively in 5.8%, 4.8% and 4.4% weight. These

values would have entireley reported in MAG fraction

except for the ultrafine sized liberated values as

explained b«low.

5.2 Effect of Size:

The size of the mineral particles have tremendous

effect on all the physical methods of separation

including separation in WHIMS. The more finer particle

experience higher drag foroe and lower magnetic force

- 324 -

and hence they have greater probabili y of being dragged

in the NMAG fraction.

The size analysis and distribution of values in

different size fractions are shown in Table VI for MURP,

SURP and RURP samples'. RURP sample has 44% of uranium

values distributed in sizes finer than 37^m compared to

28. 9% and 32% in MURP and SURP samples respectiveley.

The maximum recovery of values is only 75% with RURP

sample compared to 80-87% with SURP and MURP samples at

similar operating parameters of WHIMS. This can be

attributed to the fact that the RURP sample has higher

distribution of values in sizes finer than 37^m, and

WHIMS is less efficient In recovering from this size

rjnge. This becomes clearer from the bar-graph

(Fig.l,2&3) showing the percentage recovery of uranium

values in different size ranges. It is seen from the

graph that the unrecovered values of uranium are mostly

from the sizes finer than 37pm. The WHIMS thus seems to

be less efficient in recovering particles from this size

range due to prominence of drag forces and limitation on

magnetic force density.

5.3.Effect of Magnetic Jield and Gradient:

Both the magnetic field and its gradient are

required to be higher when a higher magnetic force

density is desired. The magnetic field is limited by

the current carrying capacity of the solenoid of the

circuit and saturation magnetisation of the soft iron

core used. The field gradient varies inversely with the

dimension of the matrix element of the WHIMS or HQMS.

The magnetic force density is lower in WHIMS compared to

HGMS because of coarser matrix elements. Thus the, lower

recovery in the RURP sample where the uranium values are

more in relativeley finer sizes compared to MURP sample

- 325 -

and SURP sample (Table VI), reflects the dominance of

drag force on those particles and report In NMAG. To

recover those fine particles of uranium minerals it may

be necessary to increase the gradient so as to Increase

the force density and this is possible with HGMS. The

results with HGMS (Table II) on classified fines of MURP

sample showed better efficiency of HGMS in recovering

fine uranium values.

5.4.Effect of Pulp Density of Feed:

The higher pulp density of feed allows more solids

per unit volume of the slurry. When the same slurry is

spread on the matrix surface, more number of particles

of magnetic material try to compete for getting captured

per unit area of the matrix resulting in entrainment of

unwanted material on the matrix surface. This

ultimateley results in poor separation and higher weight

collection in MAG. This has been observed in an

experiment with MURP sample (Table I code WH/2) where

the pulp density was 30% solids and all other parameters

aimed for maximum recovery, the weight collection in the

MAG fraction increased considerably to about 63% without

any increase in the recovery of uranium values.

5.5 Effect of Flow Rate of Feed Slurry:

This is one of the nost significant parameters in

WHIMS and HGMS for recovery of very fine particles. The

drag force on these fine particles is higher if flow

velocity is higher. This differential action of

magnetic force and drag force on a small paramagnetic

particle get accentuated by the fact that in WHIMS the

flow of slurry is in a direction perpendicular to the

magnetic field . Finer particles having relativeley

- 326 -

Lower magnetic force acting on them due to smaller

volume, therefore, are dragged in the direction of flow

if flow velocity is high.

5.6 Matrix Drum Speed(RPM):

This can be varied upto 4 rpm in the pilot plant

WHIMS. Increasing the rpm of the matrix drum results in

higher weight collection and recovery of uranium value

in MAG. It is seen from Table I (code WH/2) where rpm

was maximum resulted in higher weight collection in MAG.

The pulp density, flow rate, matrix drum speed and

magnetic field and its gradient have interactive effect

on recovery of uranium values and need to be optimized

for be3t performance.

6. CONCLUSIONS:

1. The uranium values associated with copper plant

tailings of Singhbhum are amenable to magnetic

separation and recovery of 80-85% should be possible by

WHIMS . To obtain this higher recovery of values the

weight collection of 50-55% with SORP and MURP samples

is unavoidable. RURP feed sample having higher

distribution in finer sizes results in lower recovery of

uranium values in MAG of WHIMS. The magnetic induction

required for this performance of WHIMS need to be

1.5-1.8 Tesla. The lower magnetic flux density reduces

recovery of uranium' values without reduction in weight

collection.

2. The values lost to the NMAG fractions are mostly in

the sizes finer than 37/jm. This is due to the

limitation of WHIMS in providing enough magnetic force

on these particles to overcome the dominance of drag

forces which drive them in NMAG stream.

3. HGMS can provide much higher magnetic force density

- 327 -

than WHIMS and the values lost in NMAG from sizes finer

than 37pm could be recovered in MAG in HGMS.

4. The studies show that the weakly paramagnetic

character of uranium minerals can be U3ed to recover

uranium values from ores. Eventhough the application of

magnetic separation for recovery of by-product uranium

values from the copper plant tailings may not look

attractive because of high weight collection In magnetic

fraction, its application in other ores where the

content of paramagnetic minerals like mica is low looks

definitely feasible.

ACKNOWLEDGEMENTS:

The authors thank Shri. P.R.Roy, Director,

Materials Group, BARC for hi3 keen interest in the

programme of study. They also thank UCIL for supply of

samples of copper plant tailings for study.

REFERENCES:

(1) Singh, H., Natarajan, R., Das, K.K., Jha, R.S., Sridhar,0., Rao, N.K.,and Rao.G.V.U., "Process Engg. Analysis ofUranium recovery from Copper Plant Tailings by wettabling" ,BARC/I-771,1983.

(2) Jha, R.S., Singh, H., Natarajan, R., Rambau, Ch.,andRao, G.V.O., "Investigation on Gravty beneficiation ofUranium Fines", Int. Conf. on Recent Dev. in Met. Res;Fund. 4 App. aspects, IIT Kanpur, Feb. 1985, Proc. VolPP 23-29.

(3) Corran, I.J.,"A Development in the Application of WetHigh Intensity Magnetic Separator", In " Fine ParticleProcessing" vol II (Ed. P.Somasunderan) AIME, 1980, pp1294-1309.

(4) Nesset, J.E and Finch, J.A, "Loading equation for HighGradient Magnetic Separator and Application inIdentifying the fine size limit recovery", ibid., pp1217-1241.

(5) Zimmels. Y., Lin, I.J., and Yaniv, I.," Advances inApplication of Magnetic and Electric Techniques forSeparation of fine particles", ibid., pp 1155-1177.

- 328 -

(6) Gerber, Richard., and Bris3, Robert R.. "High GradientMagnetic separator", Research studies press, John WileyfcSons Ltd,1983.

(7) Jain. S.K.."Ore Processing". Oxford IBH Publishing CoPvt Ltd. New Delhi. 1986. p 352.

Table I RESULTS WITH MURP SAMPLES

Code

WH/2WH/3

WH/5

WH/7

WH/9

MAG+MID

Xwt

62.953.1

51.1

51.3

70.2

U3°8ppm

150180

158

160

130

%R

86.387.1

80.5

80.6

94.1

ER

1.41.6

1.6

1.6

1.3

NMAG

%wt

37.147.0

48.9

48.7

29.8

°3°8ppm

4030

40

40

20*

XR

13.712.9

19.5

19.4

5.9

OperatingParameters(T.rpm.Xsd,lpm)

1.8. 4. 30. 131.8. 2, 10. 12

1.5. 2. 20. 14

1.5. 3. 15. 14

1.0. 3, 20, 131.8. 3, 20. 13

Table II: HGMS TEST RESULTS WITH CLASSIFIED FINES(MURP)

ExptNo,.

1

2

3

4

1.0

xwt

59.33

53.88

34.36

43.32

Tesla

U3°8(PP«)

162210

239

204

HAGXD

88.688.7

75.2

87.2

1.5

xwt

65.16

59.65

46.63

49.49

Tesla

Vfl(ppn)

154199207

180

XD

92.693.0

89.3

93.5

Feed Grade(J3O8(ppa)

108128

109

95

- 329 -

Table III: RESULTS WITH SURP SAMPLES

Code

WH/2

WH/3

WH/4

WH/5

WH/6

WH/10

MAG+MID

Xvrt

56.4

53.8

49.6

45.8

51.4

65.0

U3°8PPm

208

209

194

206

234

185

%R

87.3

85.3

82.7

81.3

85.8

93.0

ER

1.5

1.6

1.7

1.8

1.7

1.4

NMAG

%wt

43.6

46.2

50.4

54.2

48.6

35.0

U3°8ppm

40

42

40

40

41

26*

%R

13.0

14.7

17.3

18.7

14.2

7.0

OperatingParameters(T.rpm.Xsd.lpm)

1.8. 3. 20. 12

1.8. 2, 20'. 12

1.8, 2. 20, 14

1.8. 2. 20, 14

1.8. 2. 15. 14

1.0, 3, 20. 131.8, 3, 20. 13

Table IV-RESULTS WITH RURP SAMPLES

Code

WH/6

WH/7

WH/8

WH/9

WH/10

MAG+MID

%wt

38.2

41.3

36.6

46.8

54.7

U3°8ppm

140

134

142

164

130

\R

65.5

65.9

65.1

75.4

77.7

ER

1.7

1.6

1.8

1.6

1.4

t

Xwt

61 8

58.8

63.4

53.2

45.3

4MAG

U3°8ppm

45

48

44

47

45*

•

%R

34.5

34.1

34.9

24.6

22.3

OperatingParameters(T,rpm.Xsd,lpm)

1.5, 2. 20, 15

1.3, 4. 20. 15

1.3. 2, 20, 15

1.8, 3. 15. 15

1.0. 3, 20, 131.8. 3. 20. 13

* Final reject (NMAG) of two stage operation in WHIMS

Table V: LIBERATION CHARACHTERISTICS

xwtQuartsatlc U~0o(ppm)

X DistxwtMicaceous UoOg(ppm)

X DistLiberated U XWt6V with UgOgCppm)

Magnetite etc. X Distxwt

FEED U308rppm)

X Dist

MURP

55.817

8.338.4125

42.8

5.8940

48.9100112

100

SURP

63.618

8.831.6107

26.7

4.81707

64.5100127^

100

RURP

71.512

9.724.183

22.1

4.41395

68.210090

100

Table VI: Feed Size Analysis & Uranium Distribution

ParticleSlze(Mm)

• 208

+147

• 105

+74

+ 53

+ 37

-37

MURP

xwt11.6

23.0

21.0

15.0

10.6

5.6

13.2

%D

8.8

13.018.0

13.8

10.0

7.5

28.9

SURPXWt XD

17.4

12.5

17.9

18.0

13.6

6.3

14.3

13.2

7.0

18.2

7.5

13.6

8.5

32.0

RURPXWt XD

5.4

12.5

19.0

17.7

17.1

10.6

17.7

2.5

6.4

9.8

11.7

12.7

12.9

44.0

- 331 -

Fig 1. O I S T H I & U T I O N SC KECOVERV IN DIFFERENT SIZES

2* -

22 -

1C -

• • -

X

4 -

2 -

m

i- J 7

i^I i

•»-83 -1-74 •H47 1-208

Fig. 2. DISTRIBUTION & RECOVERY IN DIFFERENT SIZES/

-3T +3T -»03 t-74 t l O l -H47 •••20*

Fig. 3. DISTRIBUTION & RECOVERY IN DIFFERENT SIZEStim OME : msm/mt/»

+ 7* +IOI t>l47 +2Q«__CA<nicu

PRELIMINARY BENEFICIATION STUDIES ON URANIUM ORE FROM

TUMMALAPALLE, ANDHRA PRADESH

N.P.H.Padm&nabhan, U.Sridhar and N.K.Rao

Ore Dressing Section, Bhabha Atomic Research Centre.

Hyderabad.

Preliminary beneficiation studies were carried outon a small bore-hole uranium ore sample fromTummalapalle. Cuddappah district (Andhra Pradesh). Mostof the uranium values occur in this reasonably vastdeposit in the form of fine grained pitchblende. Thehost rock ic essentially dolomitic/ phosphaticlime-stone with small amounts of quartz and shale. Thepresence of such high amounts as 60-65% by weight ofacid-consuming carbonate minerals forbids the adoptionof the conventional acid-leaching process for uraniumextraction. However, if the acid consuming material inthe ore is either removed, or at the best reduced byphysical seperation method, with out any significantloss of uranium values, the acid leaching process mightstill be viable both technically and economically. Withthis aim,preliminary studies were conducted to separateessentially the carbonates by physical separationtechniques.

The ore sample contained 60-65% by weight ofcarbonate minerals, 10% of apatite and quartz each andabout 5 % of pyrite. Radiometric estimations gave theuranium assay as 0.05% U 0 *q. The ore sample wascalcined at about 900°C 3 8 for two hours and thecalcine was quenched in cold water. The slaked limeformed, is then removed by any one of the methods suchas desliming, flotation and dissolution. About 20% ofthe weight was lost during calcination by the expulsionof carbon dioxide.In the desliming method additional35% of the weight could be discarded with only about30% of loss of uranium values. In the calcination andflotation experiment, weight loss was only 20% sinceslaked lime did not float well. Attempts were made to

- 333 -

dissolve the slaked lime, and thi3 way about 70% of theweight could be discarded, with about 20-30% uraniumloss. Straight flotation of carbonate minerals withsodium oleate also, gave encouraging results. Otheralternate methods like selective dispersion of calciteor selective flocculation of apatite, quartz and pyriteare also available for the solution of problem in hand.

I INTRODUCTION

Occurrence of an extensive uranium ore deposit has

been reported by Atomic Minerals Division (AMD) at

Tummalepalle, District Cuddappah, Andhra Pradesh.*1*

The ore contains essentially dolomitic/phosphatic

limestone as the major gangue, which are highly

acid-consuming by nature, end hence, acid-leach process

for dissolution of uranium values would prove to be

unduly uneconomical, in addition to posing a major

threat to ecology and environment. Under these

conditions, alkaline leaching would naturally be a

favourable choice, but this also has its own

limitations and requirements like higher temperature

and pressure for leaching, longer contact time etc.

More importantly, Indian experience on recovery of

uranium using alkaline leach process is practically

nil, whereas considerable experience of about 20 years

exists in the case of uranium production by the

acid-leach process. Both the acid leach and the

alkaline leach routes are being thoroughly investigated

by other laboratories. There exists a third

alternatLx'G, which attempts to remove the

acid-consuming materials from the ore by suitable

beneficiation techniques in order that the remaining

ore material may be processed by the well-proven

acid-leach process. With this iaim, preliminary

- 334 -

beneficiation studies were carried out, on small amounts

of drill-core ore samples of Tummalepalle uranium ore.

Various processes were tried, in order to remove the

carbonate-bearing gangue material, and this paper

describes the experiments carried out and the results

obtained.

II CHARACTERISTICS OF ORE SAMPLE

Two drill core ore samples, weighing about 2 and

lkg were received from Uranium Extraction Diviaion(UED)

of BARC for beneficiation studies. Since the quantity

of the ore sample was so small, batch beneficiation

experiments were carried out with about 50-100gm of

feed per batch test. The experiments are, therefore,

only exploratory, and the results only indicative.

Detailed studies could not be continued due to the

paucity of the ore sample. However, these experiments

do indicate that prior to the outright rejection of the

proposal to set up a uranium extraction- plant for

processing this ore on techno-economic grounds, this

technique of removal of acid-consuming material from

the ore can be given a serious consideration.

Mineralogical analysis of the ore samples

indicated that this ore contained 60-65% by weight of

carbonate-bearing minerals, 10% of apatite and quartz

each and about 5% of pyrite. Radiometric assay gave

the uranium assay as 0.05% UB0B#<». The main uranium

bearing mineral was found to be pitchblende, while

minor amounts of coffinite was also reported. Apatite

also showed a small amount of radioactivity. The

distribution of particle size and uranium values in a

TABLE I

DISTRIBUTION OF URANIUM VALUES AS A

FUNCTION OF PARTICLE SIZE

SIZE FRACTION

MESH

+ 35

- 35 + 50

- 50 + 70

- 70 + 100

- 100 + 150

- 150 + 200

- 200 + 270

- 270

FEED

WEIGHT

%

26.9

17.3

9.1

11.7

5.6

4.1

3.0

22.3

100.0

* °3°8ASSAY

0.048

0.045

0.046

0.043

0.041

0.041

0.035

0.035

0.043

U3O8DISTN.

%

30.2

18.3

9.7

11.7

5.4

3.9

2.5

18.3

100.0

crushed samplefTable I) showed that there is no

preferential concentration in any size fraction.

However, it may be noted that the coarsest size

fraction (+35 mesh) has slightly higher uranium assay,

while the finest size fractions (-200+270 and -270

mesh) have slightly lower assays. The high uranium

distribution in these sizes are mainly due to the high

weights of these si2e fractions.

Ill BENEFICIATION STUDIES

The problem of rejecting the carbonate-bearing

gangue is generally encountered in the beneficiation of

rock phosphates, and a variety of processes are

commonly practised all over the world. Flotation and

thermal methods are important among them. The idea of

- 336 -

thermal methods is not new in processing of uranium(2)ores also. One of the uranium plants in Western

Australia (Western Mining Corporation Limited, at

Yeelirrie) is rejecting the dolomite and calcite

present in the ore by roasting in a rotary kiln,

followed by quenching. In the present study also,

experiments were carried out based on thermal treatment

and flotation. In the case of thermal experiments, the

calcination and quenching were done under similar

conditions, but subsequent operation differed from

experiment to experiment. Essentially the experiments

wore carried out as mentioned below •

(1) Calclnation-Quenching-Desliming

(2) Calcination-Quenching-Flotation

(3) Calcination-Quenching-Dissolution

(4) Direct Flotation

3.1. Calcination - Quenching - Desliming

The ore crushed to,all passing through 6mm was

calcined in a laboratory muffle furnace at 960°C for

2 hours, and the calcine was quenched in cold water.

Calcination results in the expulsion of carbon dioxide

from the ore, and quenching oauses thermal stress in

the quick lime, because of which the material gets

fragmented and forms slaked lime in water. The

temperature and duration for calcination were arrived

at based on our earlier experience on the beneflclation

of Maton Phosphate Ore. The material was then

deslimed in a controlled manner to remove the slaked

lime in the overflow. The results are presented in

Table II. (Expt.l). About 25X of the weight was lost

- 337 -

during calcination, by the expulsion of carbon dioxide.

In the desliming stage additional 35% could be

discarded with about 30X loss of uranium values. The

uranium assay went upto 0.1% ^90m. which was

incidental. The slaked lime was found to flocculate at

high pH ( the high pH being due to the presence of

lime) and hence, desliming was less efficient. Attempts

were made to prevent flocculation of the slaked lime by

carrying out the quenching operation in

dispersant-mixed cold water instead of plain water.

TABLE II

RESULTS OF CALCINATION-QUENCHING DESLIMING EXPERIMENTS

FRACTION

Sands

Slimes

Wt.Loss

Feed

Expt.1

WT X

41.3

35.0

23.7

100.0

AssayXU3O8

0.1

0.05

0.059

Dlstn.U3O8 %

70.2

29.8

100.0

Expt.2

WT X

35.8

39.3

24.9

100.0

AssayVJ3O8

0.12

0.04

0.059

Dlstn.U308 X

73.2

26.8

iOO.O

Sodium silicate was used for this purpose in one of the

experiment and the remits are given in Table II under

Expt.2. Sodium silicate was not effective both in the

prevention of flocculation of the slaked line and

dispersing the already flocculated lime. More efficient

dispers^nts, like sodium hexa metaphosphate or low

molecular weight polyacrylates are expected to give

better results. The relatively high loss of uranium

values could mainly be due to the inproper dispersion

and desliming.

- 358 -

3.2. Calcination - Quenching - Flotation

After calcination and quenching, flotation was

tried to remove the slaked lime particles in the float.

Sodium oleate was used as collector and methyl isobutyl

carbinol (MIBC) as frother. The results are given in

Table III. There was a weight loss of 21% during

calcination and quenching and an additional weight of

about 20X could be discarded during flotation, with

about 16% loss of uranium values. Slaked lime did not

exhibit good flotation properties, and hence the weight

TABLE III

RESULTS OF CALCINATION-QUENCHING-FLOTATION EXPERIMENT

FRACTION

Float

Tails

Wt.Loss

Feed

Expt.3

WT %

19.6

59.1

21.3

100.0

AssayXU3O8

0.047

0.08

0.056

Dlstn.U3O8 %

16.3

83.7

100.0

loss during flotation was found to be less than

expected. Similarly, with improved dispersion it should

be possible to minimize uranium loss in the float.

3.3. Calcination - Quenching - Dissolution

In the beneficiation of phosphates, the slaked

lime is removed in some of the plants by dissolution,

followed by solid-liquid separation. The dissolution is

achieved by increasing the solubility of lime with the

- 359 -

help of reagents like ammonium or sodium chloride.

The solubility of slaked lime is claimed to increase by

abcut one hundred times if sugar solution is used for

dissolution. ' The lime reacts with the sugar

molecule as follows [S(OH)2 indicates sugar molecule] :

/ OH . 0 .( + Ca(OH) • S ( ) Ca + 2H 0\ OH 2 \ o / *

The calcium complex of sugar is soluble in water, and

can therefore be removed by simple solid-liquid

separation.

Indicative experiments were carried out on the

Tummalepalle uranium ore sample, using this technique.

Calcination was carried out at 880°C for one and a half

hours as recommended by Gunduz and Guagum. Quenching

was done in cold water, and after removing the excess

water, strong sugar solution was added, and the mixture

was stirred for about 30 minutes. Then the solids were

removed by decantatlon and the supernatent, liquor was

filtered, to get fines and filtrate. The filtrate did

not show presence of any uranium values. The

results, given In Table IV, show that the total weight

loss that could be achieved by calcination,

quenching and dissolution was.about 43%, and the solids

weighed 36X, containing 81% of the uranium values,

while the fines weighed 20.7%, and contained 19% of

uranium values. Since the filtrate does not contain any

uranium values, and since there is no other product

wherein uranium could get distributed, the solids and

fines together constitute 57% by weight, contain 100%

of uranium values, and the assay of the combined

material (i.e.. solids and fines) will be 0.095% Ua0,.

- 340 -

TABLE IV

RESULTS OF CALCINATION-QUENCHING-DISSOLUTION EXPERIMENT

FRACTION

Solids

Fines

Wt.Loss

Feed

Solids

Fines

Expt.4

WT %

36.5

20.7

42.8

100.0

57.2

AssayXU3O8

0.12

0.05

0.054

0.095

Distn.U3O8 %

80.9

19.1

100.0

100.0

Another experiment was conducted on similar lines, but

with the addition of deslimlng step before dissolution.

But this experiment did not give good results, as in

the earlier cases, due to improper dispersion of slaked

llae and sub-optimal desliming. About 33% of uranium

values were lost in slimes during the desliming stage.

However, it is felt that the results would be better if

the desliming step is introduced after dissolution, as

this would facilitate the subsequent solid-liquid

separation, and with acceptable loss of uranium values

the weight could be reduced by an additional 20%.

3.4. Direct Flotation

One exploratory experiment was carried out using

direct flotation to discard the acid-consuming

carbonate gangue in the float. Flotation was carried

out on a ground ore sample with sodium oleate as

collector and MIBC as frother, at a pH greater than 10.

- 341 -

TABLE V

RESULTS OF DIRECT FLOTATION EXPERIMENT

FRACTION

Float

Tails

Feed

WT

61

38

100

X

.8

.2

.0

Expt.5

AssayX03O8

0.03

0.12

0.060

Distn.U3O8 %

30

69

100

6

4

0

The results (Table V) are good in terms of the removal

of gangue and weight reduction, but not so good in

terms of the accompanying loss of uranium values.

About 62% of the weight could be discarded in the

float with about 30% of the uranium values. It needs to

be mentioned here that a two-stage flotation process

has been developed for the phospho-uraniferous ore of

Itataia in Brazil. ' The ore is calcareous

phosphorite, with uranium occurring to the extent of

0.116% UB0#. The stragtegy adopted in processing this

ore Involves direct flotation of calcite and apatite in

the first stage, using a reagent combination including

sodlun silicate, starch, sodium hydroxide and tall oil

at a pH of 10, and a reverse flotation in the second

stage, with depression of apatite using phosphoric acid

and activation of calcite using sodium oleate at a pH

of 5.5. Uranium reports along with apatite and their

pilot plant tests Indicate that 63.6% of uranium values

could be recovered in the apatite concentrate with a

grade of 0.204% U-0#. About 15.4% of the uranium values

are lost in the slimes (<10JJR) while about 17.6% report

in the tailings of the first flotation stage. The

flow-sheet is given in Figure 1, along with material

- 342 -

KEY WT %U3O8 AssayXU3O8 Dlst.%

100.00.115100.0

Ground Ore Feed

IDESLIMING CYCLONE

81.70.11984.6

Under

SilicateTails

DIRECT

flow Overflow

LOTATIONPH = 10

32.20.0617.6

Float

180.15

<L.309.4

REVERSEPH

Sink

360.63

FLOTATION= 5.5

j.0204.6

CalciteFloat 1

13.50.0293.4

Apatite Concentrate

Figure 1. Flow-Sheet for ProcessingPhospho-Uraniferous Ore

Itataia

balance for uranium. A similar process strategy could

be thoroughly investigated in the case of Tummalepalle

ore also. Since uranium is of Interest in the present

case, the silicate tailings (sink of first stage) and

apatite concentrate (Float of second stage) could be

nixed to get higher uranium recovery and attempts can

be made to recover the uranium values lost In slimes,

using other techniques like high gradient magnetic

separation (HGMS). A similar two-stage flotation

process has been recommended for removal of dolomitlc

impurities from Jamarkotra phosphorites. '

- 343 -

IV DISCUSSION

As already mentioned, all the above process

schemes could not be studied in detail due to want of

ore sample. However,the preliminary experiments

indicate that it is quite possible to acheive the main

objective of removal of acid-consuming materials prior

to leaching of the ore, and that an in-depth study is

essential for evaluating the technical feasibility and

economic viability of the whole scheme. Among the

various processes tried, calcination- quenching-

dlssolution and direct flotation processes appear more

promising. In addition to these, other processes like

selective flocculntion and selective disi-jrsion

followed by suitable separation techniques, high

gradient magnetic separation etc. can be explored.

V CONCLUSIONS

A few exploratory experiments were carried out on

a small amount of bore hole ore samples from

Tummalepallc. Andhra Pradesh, with the objective of

reducing the acid-consuming carbonate-bearing gangue

material, by thermal methods and flotation. The

preliminary experiments indicate that it is possible to

achieve the objective with minimum uranium loss. About

50-60% of the total weight could be discarded with a

uranium loss of 20-25%. Although the primary aim was

not to upgrade the uranium content in the ore, the

f.rade of th* ore goes up to more than 0.1% Us0a from

0.05%. Among the various processes tried,

calcination-quenching-dissolution and direct flotation

- 344 -

processes give good results, and can be taken up for

serious studies.

Acknowledgements

The authors would like to express their thanks to

Shri.P.R.Roy, Director, Materials Group, BARC., for his

keen interest in the problem and to Shri.K.S.Koppikar,

Head, Uranium Extraction Division, for the supply of

the ore samples.

VI REFERENCES

1. Annual Reports of Department of Atomic Energy,

1977-1988.

2. Significance of Mineralogy in the Development of

Flow-Sheetfor Processing Uranium Ores. Techniocal

Report series 196, International atomic Energy Agency,

1980, P45.

3. - ibid - , p264.

4. Rambabu Ch., Roy P.K.. Shukla S.K., Majumdar K.K.

and Rao G.V.U., Beneficiation Studies on Maton

Phosphate Rock (Rajasthan), BARC Internal Report

BARC/I-857, (1978).

5. Ben-Ari C. and Fuchs W.J., Upgrading Calcareous

Rock Phosphate, Neger Phosphates Limited, 1960.

6. Herman E.R., Upgrading of Phosphate Rock, Chemicals

and Phosphates Limited, 1965.

- 545 -

7. Kirk K.E. and Othmer D.F. (eds.) Encyclopedia of

Chemical Technology, vol 12, Wiley Inter Science, New

York, 1967.

8. Gunduz T. and Gumgura B., The Enrichment of

Low-Grade Mazidagi Phosphates by Calcination and

Extraction Methods, Separation Science and Technology,

22(6), pp 1645-1648, 1987.

9. Acquino J.A., Furtado J.R.V. and Reis(Jr.) J.B.,

Concentration of Phosphate Ore with

Siliceous-Carbonated Gangue via Reverse Flotation,

FROTH FLOTATION : Proceed. 2nd. Latin-American Congr.

Froth Flotation., Concepci6n, Chile, 1985

Developments in Mineral Processing, vol 9, ed: Castro

S.H. and Alvarez J., Elsevier (1988), pp 185-200.

10. Moudgil B.M. and Ince D., Flotation of Dolomite

Impurities from Jamarkotra (India) Phosphorites, Inter.

J. miner.Process., 24(1988), pp 47-54.

- 346 -

Session III

DI5CU3SI0HS

Paper No. 1

N. SV/AMINATHAR s What is the role of blasting technique withrespect to ore dilution?

J.L. BHASIN : Supervision of blasting operation is veryimportant. Dilution of ore to the extent of about 10 per centis tolerable.

R. MOHANTY t What is the difference in mining cost with respect

to depth ?

J.L. BHASIN : As you go deep the time taken to bring out agiven quantity of ore Increases reducing the productioncapacity of the mine and hence the mining costs increase withdepth*

Paper Ho. ?

K.K. DY/IVEDY t What la the additional recovery by using BMS unit?

U.K. TIWARI t We have not been eble to increase the recovery inour plant trials beyond additional 8-10 per cent using Mosabonltailings.

N.K. RAO t Are you using the aame type of table in all yourplants or are you using different types?

U.K. TIWARI i More or less same type of tables are used. Ofcourse some modification in the tables have been made from time totime based on the operating experience* *

- 547 -

N.K. RAO* What waa the reason for the failure of KDCC cone,concentrator in upgrading uranium from copper tailings?

U.K. TIWARI: This could be due to the association of uraniummineral with fines and the fine grind of feed material*

N.K. RAO : Is it advisable to use the tables first and then BMS?

U.K. TIWAHI » We have tried this type of flow sheet. Of courseBMS having very high capcity could be the choice for the firststage.

M.C. BHURAT > Why not apply direct leaching technique for coppertailing?

K.3. KOFPIKER t I would like make comment. In the next sessiona separate paper is being presented on direct leaching of uraniumfrom copper tailing. Hence this aspect can be discussed at thattime*

S. SEN t TV» y O U cave any future programme to improve the recovery?

U.K. TIWARI t Efforts are made on continuous basis.

Paper Ho. 4

N. SWAMINATHAN » Do you have any control on the particle sice

of the tailings?

R. 3HANKARAN t I would like to add my comment* We do not have anycontrol on particle else beoause grinding Is done by H.C.L. to salttheir copper recovery circuits.

K.K. DWIVSDY t I feel that either magnetic separation or direotleaching should be followed. By magnetic separation aboutof uranium can be reoovered and them it be leached.

- 348 -

Paper No. 6

D.V. BIIATNAGAH : Magnetic separation may give a better recovery

at the concentration stage. One should test this concentrate

to determine how much of uranium present can be leached.

!?.?. V5RMA : What would be the average grade of this concentrate?

K.K. DWIVEDY t I would like give some data. In our laboratory

studies, the concentrate assayed 0.022$ U,Og and leachability

T.K.S. HUETHY i I would like make some general remarks. The

work on the recovery of uranium from copper tailings has been

going on fov the past three decades* I think the time has come

when all the parties involved sit together and reach some

concrete decision.

Paper No* 7\

X.K. DWIVEDY t I would like make a comment. For this type of

ore the only solution is to float the sulphide and then go for

alkaline leaching because In carbonate flotation, it is difficult

to get rid of total carbonate*

D.V. BHATNAGAR : I suggest that it is better to remove the

sulphides and follow alkaline leaching technique.

I). 0. BANNKHJKB : The apatite content of the ore from this area is

not uniformly 10 per cent. It varies widely. We should plan

on the lines suggested by Shri Bhatnagar and Shrl Dwivedy*

S E S S I O N I I I B

ANALYTICAL TECHNIQUE'S IN URANITTC TECHNOLOGY - I I

Chairman : Shri L.M. MAHAJAN(Retd. )

i? A R C

Reporteurt Shri N. S^AM SUNDARN ? C

- 349 -

RAPID DETERMINATION OF URANIUM IN URANYL NITRATE

SOLUTIONS BY GAMMA SPECTROMETRY

T.K. SANKARANARAYANAN AND D.S. GUPTA

Chemical Engineering Division

S.G. SAHASRABUDHE AND M.R. IYER

Health Physics Division

And

V.N. KRISHNAN

Uranium Extraction Division

SUMMARY

Uraniuo-235 emits gamma rays of 185.7 kev with 54Z yield. Gamma

spectrooetry with this gamma ray presents an excellent method for the

estimation of U23S and hence the total natural uranium concentration.

This paper describes the setting up of a system which uses a 3"x2" Nal

detector and a microprocessor based 4K multichannel analyser for

assaying U in liquid samples, A software gain correction method is

Incorporated in the system to eliminate errors due to gain shift.

Three different ranges of concentrations of Uranium in natural Uranyl

Nitrate solution have been studied. Using known, standard samples,

calibration graphs of cps Vs. concentration were obtained. Linearity

has been observed upto the Uranium concentration of about 80 gas/litre

while non-linearity due to self-absorption was found in the higher

ranges. An exponential relationship* vix&x C. wa* used for fitting the

data in the non-linear range.

This method can be used to determine the Uranium concentration In all

the three ranges of concentration we have studied in a rapid and

non-destructive way. The standard deviation in the concentration range

of 50-80 gms. of U per litre was found to be + IX for a 300 sees.

counting time.

- 350 -

1. INTRODUCTION

Uranium-235 emits gamma rays of 185.7 kev with 54Z yield accompanying

the alpha decay of U-235 to Th-231. This gamma line from U-235 detected

by a Nal(Tl) detector has been used for precision online measurement of

Uranium enrichment in a LWR fuel fabrication plant (1) . Gamma

spectrometry using 185.7 kev gamma ray presents an excellent method for

the estimation of U-235 and hence the total uranium concentration in

natural uranium samples. A method based on the above technique has been

developed for the rapid determination of uranium in uranyl nitrate

solutions non-destructively. The concentration ranges studied are those

normally encountered in uranium refining plants.

2. PRINCIPLE OF THE METHOD

Assuming the isotopic composition of uranium to be natural and uniform

in all samples of uranyl nitrate solutions, the intensity of the 185 kev

gamma emitted will be proportional to the U-235 content which in turn

will be proportional to the total uranium content in the sample. If the

sample geometry, volume of the sample and the matrix are kept the same

in the standard and the samples, the calibration graph obtained with cps

vs. concentration of U in gms/litre =an be utilized to determine the

concentration of U in samples.

3. EQUIPMENT

A microprocessor based HPD 4K multichannel analyser was utilised for

this work. The block diagram of the equipment is given in Plg.l. A Hal

(Tl) detector (size: 5 cms. x 7.5 ems) coupled to a phot'omultipller was

used along with a preamplifier, spectroscopy amplifier, ADC and 4K MCA.

Cylindrical PVC jars of dia. 9cms, haying a gasket and a lid were made

use of as sample bottles. In order to reduce the systematic errors due

to geometry effects the sample container was positioned on top of the

detector using a guide ring.

- 351 -

4. PROCEDURE

Using PVC jars as sample containers, solutions of uranyl nitrate were

kept on the Nal detector and the gamma spectrum acquired for a counting

time of 300 sees. A typical spectrum obtained from the samples is given

in Fig.2. A peak at 90.7 kev due to X-rays and decay gammas Is seen in

addition to the 185.7 kev gamma peak from U-235. The method involved

the estimation of the photopeak area of the 185kev peak after correcting

for the contribution from Compton scattered gammas of higher energies.

A Conpton window adjacent to the 185 kev peak on the right side is used

for finding the Compton contribution using a linear Compton

approximation. The selection of the peak and Compton window would be

subject to personal errors and the spectrometer settings like gain etc.

A software controlled gain correction method developed and incorporated

in the MCA avoids personal errors and ensures that any gain shift

resulting in peak shift in the spectrum is automatically taken into

account for arriving at the photopeak area. The procedure enables the

photopeak area to be estimated within an accuracy of better than lit.

The method consists of Internally arriving at the energy calibration

using 90.7 kev peak and 185.7 kev peak in the sample spectrum itself.

The two windows for the 185.7 kev photopeak and for the Coapton

contribution to the peak Incorporated in the software of the analyser

are specified In energy units:

185 kev peak window: i. 156.13. - 236.56 Jeev

Coapton window : 241.68 - 275.19 kev

Using the Internal energy calibration the corresponding channel windows

are Internally arrived at and the Integrated counts In the windows are

arrived at by the analyser. A 300 sees, counting of the sample was

found to give sufficient counting statistics to ensure better.than IX

overall precision for concentrations above 50g U/l . The optimization of

sample volume was carried out. A volume of 200 ml. of sample was found

to be optimum on the basis of these studies.

- 352 -

5. RESULTS

Three different ranges of concentrations of uranium in pure uranyl

nitrate solution were studied. The ranges selected are those of

interest in various types of samples in a Uranium Metal Plant.

Range 1: 0.200g U/l to 1.4 g U/l.

When background effects were completely eliminated by proper shielding

of the sample, a linear graph passing through the origin was obtained

for cps Vs. concentration of U in gins/I it re. A linear regression was

carried out on this data which resulted in the following expression:

g/1 of U - 0.1866 x cps + 0.0106

The fitted values agreed with the actual values of concn. within 2.4%

Range li: 50 gms U/l to 80 gas.U/l

Linearity was observed in this range also, when cps was plotted against

the concn. of U In gas/I which is shown in Pig.3. Linear regression was

carried out which yielded the following expression:

gms/1 U - 0.180 x cps - 15.48

The maximum deviation of fitted values from the actual values was only 0.4%

Range ill: 150 gms U/l to 190 gms U/l.

Due to self absorption of gamma rays by the solution, non-linearity was

observed In the calibration graph of cps. Vs concn. of U as seen In

Fig.4. The following equation was fitted using the data:

-Cxy - Bxe ; where y Is the cps obtained for a given concentration

x. Typical value for B and C are 5.808 and 0.00205

respectively.

6. DISCUSSION

When raffinate solution was counted, the radium present Interfered with

- 353 -

the 185.7 kev gamma rays, ^hereby resulting In an erroneous value.

Radium has an emission at 186 kev and under equilibrium conditions, 50%

of the contribution to 185 kev photopeak comes from Ra-226. Generally,

Ra-226 Is not present to significant levels in the uranium handled in

Uranium plants engaged in processing of nuclear fuels since it gets

removed in the earlier purification stages. However, in the raffinate

stream, Radium-226 may preferentially get collected and could pose a

problem in the type of measurements described above.

In such cases, the radium has to be chemically removed before the gamma

counting or alternatively the contribution from radium can be corrected

using one of the many gamma emissions from Its daughter products which

will always be in equilibrium with Ra-226 such as the 352 kev emission

from Pb-214 with a branching ratio of 37.It.

The results reported above indicate that this method is Ideally suited

for a rapid non-destructive assay of pure Uranyl nitrate solutions of

concentration range 50-80 gms./l. Even for higher uranium concentration

the method can be applied using an exponential expression with

empirically determined values for the constants. The accuracy of the

method In the linear region is found to be +_ IX.

ACKNOWLEDGEMENTS

The authors are grateful to Shrl T.K.S. Murthy for suggesting this

problem and fruitful discussions and also to Shri V.S. Keni, Head,

Process Engineering Section and Shrl S. Sen, Head, Chemical Engineering

Division for encouragement in carrying out this work.

REFERENCES

1. Gamma ray spectrometry for In-line measurements of U235 enrichment

in Nuclear Fuel Fabrication Plant (IAEA/SM/201/46); P. Matussek and H.

Ottmer; Safeguarding Nuclear Materials; Proc. Symp.; Vol.11, pp.223.

- 354 -

PHOTO-MULTIPLIER

PREAMPLIFIER

Not ITl)CRYSTAL

A.O.C. CONTROLLED4KUCA

INPUT/OUTPUTDEVICE

TOUTPUT

FlG.1. BLOCK DIAGRAM OF Y SPECTROMETRY SYSTEM FCRNP ANALYSIS OF URANIUM

J-Wt-70 Rtv.

FIG. 2 . GAMMA SPECTRUM OF URANYL NITRATE SOLUTION

•oo

8

» toMM CO • • 70 7t

3 CONCENTRATION OF U IN «••/Litre

790

ISO

Fit.4

170 190

CONCCNTRATION OF U IN

- 356 -

MODIFICATION OF FLUDRIMKTRIC MKTHDD OF DRANIUM ANALISI3

FOR JAPtPOPA PLANT SiMPLSS

AJB, Chakraborty and V.M. Pandey

riON OF INDIA

JJDOGUDA MIMES

SINGSHUI

BIHA&

Fluorimetry i s one of the most sensitive instrumental method* of est i-

mating uraniuB. The method followed at present Involves the extraction

of uranium vith ethyl acetate in presence of saturated solution of

aluminium nitrate* After extraction, an aliquot of the extract i s pipe-

tted into platinum dishes specialty made for f luorlaetric work and the

solvent i s evaporated under an Infra-red lamp. The residue i s fused with

about 0.4 gm of aodiua fluoride - sodium carbonate (1:4 mixture) at a

temperature of about 800*0 for 3 minutes using a muffle furnace. The

fused mass i s cooled and the fluorescence of the resultant bead i s

measured.

The samples analysed by f luorimetrie method In our laboratory are ( l )

Break through, (2) Semi pregnant, (3) Barren Diversion, (4) Second Duetes,

(5) Grab sample of eluate, (6) Secondary f i l t er cake, (7) Barren liquors,

(6) Leech Tailings, (9) Plant Tailings.

tfhile using ethylacetate, extractions are done in nitrate medium whereas

most of our samples are in sulphurlo acid medium* H»nce a solvent suited

for sulphate medium was fe l t to be more useful* Jnines are being used ex-

tensively to remove uranium from sulphate liquors as an anion* Alemine

336 has been used in our BAD studies for solvent extraction of uranium

from Jaduguda leach liquors* Since i t was found to be a good extractant,

the same solvent was selected for extraction for fluorlmetrio analysis

of uranium in place of ethyl acetate ti11—<"<"" nitrate. It was found that

Alamine-336 can be used in plaoe of ethyl aoetate aluminium nitrate for

- 357 -

uranium extract ion for f luor imetr ic determination with the same accuracy

as in the case of e thyl acetate aluminum n i t r a t e .

INTRODUCTION

The f luorescence of uranyl compound on i r r a d i a t i o n with Ul trav io le t

l i g h t i s wal l known e f f e c t . I t was discovered by Becquarel and Stokes

i n t h e middle of previous century, s ince that t ime the phenomenon has

been care fu l ly invest igated by many authors. One of the r e s u l t s of t h e i r

e f f o r t s was the discovery of t h e natural r a d i o a c t i v i t y and another con-

sequence vas t h e development of a very s e n s i t i v e method for t h e qua l i ta -

t i v e and quant i ta t ive determination of uranium (Uranium Fluoriaetry) in

water, minerals , b i o l o g i c a l mater ia l s e tc* Fluotfl metric method for d e -

termination of uranium in so lu t ion i s in use r i g h t f roo t h e day* of

Manhattan Project i n U.S.A. The method followed at present i s mostly i n

t h e l i n e with the one proposed by Grimaldi and further improved by Centanni,2

Bos* and Oesesa . The procedure followed at Jaduguda for process stream

samples invo lves separation of uranium from sample so lut ions using e thyl

acetate a* an extract ant i n presence of saturated solut ion of aluniniuB

n i t r a t e , drying of * measured al iquot of ex trac t , fus ing t h e dried a l l .

quot wi th sodium f luor ide - sodium carbonate f l u x and than determining

the uranium by f lnor imetr ic method. S ine* most of the plant samples are

i n sulphuric acid medium, a solvent su i tab le for sulphate medium was f a i t

t o be more use fu l i n place of e tby lacetate - Aluminium n i t r a t e system.

Therefor a i t was decided t o u s * a solvent which can extract uranium In

sulphate medium.

Long chain a l iphat i c amines are ooing used ex tens ive ly i n uranium industry

for uranium extract ion from sulphurlo acid loach l i q u o r s . One of the

popular amines, Alav.Jie-336 was used i n our B&D s tudies for uranium a t -

t r a c t i o n from Jaduguda leach l i q u o r . Since i t was found t o be a good e x -

t rac t ant for uranium i n sulphurlo acid medium, t h e saae solvent vac ,

so lsctod for extract ion of uranium i n process stream senples for f l n o r i -

a o t r i c est imations of uranium and the es t imat ions were carried out suoo-

- 358 -

easfully. Studies were farther extended for assaying niU feed samplessod sample• froa other sines of Singfrhhai in place of T.B.P, extractionspeotrophotoaetrio aethod. The results obtained are dealt with in thispaptr.

(a) Seageats and Ch—deals

( i ) Xtamlno-336 ( l* VA solution) t~ ID ml of alaalno-336 was

diluted to 1 l i t re using IR grade bensene. The diluted solventwas washed twice with 100 a l portion of water of pH 1.0 andwas filtered with whataaiw40 f i l ter paper and stored in abottle.

(11) Flux Mixture!- 4 i l Mixture of sodlua oarbonste aid sodluafluoride .

( i l l ) Standard ttraniusi Solatloni- (lO^ng/al) in sulphuric acid

•edium.

(b) aaaple Preparation*. 1 ga of the powdered ore/Tailings saaple wasdigested with 50 a l of an aold aixturc oontaining 5 a l of nitricadd, 5 a l of sulphuric aojd and rest water for 1 - 4 hours on ahot plate and evaporated to dense sulphurie acid fusing. The fusingwas continued for SO to 45 ainutee and then the beaker was oooladand SO a l of water was added earefully end was boiled for 5 to 10alnutes. ifter cooling It was filtered and the voluas was aede upto 250 a l la cav> of ores and 100 a l in cass of tailings*

(o) Uraniua &ctractlon froa Solutions*-. Ursniua was extracted withAlamlne-356, as an extract ant in benaene diluent,' The effect ofother diluents such as, cgrolohexane, ether and etbrl aoetst«f con-oentration of extraetant, pH of the aqueous phase, tiae end phaseratio on the extraction of uraalua were studied* The used solvent

- 359 -

was regenerated by washing twice with 2.5 percent sodium carbonate

8olution(l l i t r e of used solvent with 100 ml of 2.5£ sodiun carbonate

solution) followed by one wash with 1.0 pH water and f ina l ly with d i s -

t i l l e d water.

(d) Procedure for analys i s ! - A suitable aliquot of the solution

samples of the plant and of the ores and t a i l i n g s as prepared

above was taken i n stoppered extraction tube and volume was made

upto 10 ml keeping the pH i n the range of 1.0 to 1 .5 . 10 ml of one

percent solution of alanlne-336 In benzene was added and was shaken

for 5 minutes and was allowed to s e t t l e for complete phase separa-

t i o n , 0 . 1 ml or 0 .2 ml of t h e extract was pipetted out into the

f luorimetrio platinum dishes and the solvent was evaporated under

IB. lamp. 0.4 t o 0.5 ga of sodium carbonate sodium fluoride f lux

mixture was put into the dishes and was fused at 800°C for 3 minutes.

After cooling the intens i ty of fluorescence of the fused beads

were measured with the help of Jarre 1 Ash Fluoriaeter. A set of

standard* and experimental blank were run through the «*»<1aT>

procedure and the Intens i ty of fluorescence measured were oompared

t o know the unknown concentration.

The e f fec t of addition of interferences on the accuracy of uranlua e s -

timation by using Alamine-336 ware also studied. Experiments were also

conducted for supresslng the interferences.

RJBULTS AMD DISCOSSIOMS

1. Test for the Sui tab i l i ty of Jbctractant Concentrations

Uranium was extracted from the solution containing different amounts

of l^0Q using & alamine-336 i n benaene. A. plot of f luoriaetrlo

readings against quantity of UgOg in solution i s shown i n ? i g . ( l ) .

The straight l ine in F i g . ( l ) shows that upto SO ug of 1%0Q oan beextracted by Xl alamine-336.

- 360 -

(Z) Gonparislon of Uranium Estimation Involving Two Dlffere.it Matted

of

The extracted uranium from the process stream samples by using onepercent alamine-336 in benzene and by using ethyl acetate abminiuani t rate were analysed fluorimetrically and the results are comparedin Tabla-I.

TiBLE - I

Samples (ga/1)

{After extraction with j ifter extraction withIl£ alamine-336 in benzene I ethyl acetate AL-I f nitrate medium

Ion exchange Barren

Plant t o t a l BarrenBarren DiversionSecondary f i l trateBreak through sample

0.0052

0.0033

0.0100.0137

0.0013

0.00540.00320.0100.01300.00135

The results show that there i s a very good agreement in tho resultsobtained which confirms the applicability of alamine_336 as an extract antin place of ethyl acetate aluminium nitrate system to extract uraniumfrom sulphate media.

(3) Coapariaion of Ireoision of Jaslyals Using Two different Extract ants

for Plaorlmetrio Determination of Oraniun.

To compare the precision of analysis three different samples wereanalysed several t l»e by using ilsmlne-336 in bensene and by ethylacetate w"1'—<«fc»» nitrate for the extraction of uranlun. The resultsare given in Table-U.

- 361 -

TiBLS- n

DETffiMIHATIOM OF URANIUM IN THS PROCESS SAMPLE

'A1 - 3y alanino-336 in benzene extraction

•B • - By ethyl acetate aluniniiim nitrate extraction

Sample

Barren .

diversion

Secondary -pulp

filtrate

A

B

A

B

A

B

No. ofdeter,mination

7

7

5

5

7

7

Value obtainedg/1 IU>8

0.00196

0.0020

0.0310

0.03ID

0.034

0.034

0.001760.00176

0.0290

9.0285

0.031

0.031

I Average value]

0.00184 O0.00186 O

0.0299

0.0293

0.0323

0.0320

Standarddeviation

•000071•000085

0.00081

0.0010

0.00106

0.00106

3 3 S S X S S S = s s s s a

From the result* i t i s clear that the precision of analysis in case ofalanine-336 extraction i s batter than that of ethyl acetate aluainiuanitrate systea.

(4) application of ^ ine Extraction in the Analysis of Tailings

of the giant.

The tailings samples were analysed for uraniua f luoriaetricallyusing alenine-3'36 in bensene and ethyl aoetate-alaainlw nitrateas an extractant. The results obtained by two different routes

oonpared in Table-Ill.

- 36? -

- niANALYSIS Oy TAILINGS SAMPLES USING TMD DIFFgtEHT EXTRACXANTS

S a m p l e s *°3°8

Sxtractantas an I Using AL-nitrat© and Ethyl-

I Acetate as an Sxtractant

Leaching Pachuca

Tailings

Tailings pachuca

(Final Tailings)

0.0048

0.0042

0.0040

0.0040

0.0041

0.0042

0.0075

0.0077

0.0078

0.0078

0.0077

0.0084

0.0044

0.0042

0.0045

0.0042

0.0043

0.0041

0.00800.00780.00810.00760.00760.0079

S S 3 B

From the results In the table i t can be seen that the ralnes obtainedby using two different extract ants are In rery close proximity.

(5) Application of Alamlne Extraction for the Analysis of Uranlw in

Dranif Ores of Slnghbhu* Belt.

Daily samples of classifier overflow product (GOP) of Jaduguda plantand uranium ores from Narvapahar and Turandih were analyael foruraniua f luorlaMtrically using alamlne-336 in. bensene as an extractantand spectrophotoaetricalljr using T.B.P. extraction method. The resultsof OOP analysis by two different methods are oompared In Table-IVand of Harvapahar and Turamdih ores are compered in Table-V.

- 363 -

TJPLB - IV

OOMPAHISION OF COP ANALYSIS BY TWO DIFPBtHfl METHODS

Date * U 3 ° 8 I Date

By Colorine-j By fluorl-tr ic method | metric method

L

* D 3 ° 8

By Colori-metricmethod

|By f luorime-[tric method

1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

0.0600.059

0.058

-

-

0.062

0.C59

0.0600.057

0.058

0.0600.0600.085

0.061

0.057

0.0610.058

0.060

-

-

0.062

0.058

0.C600.C56

0.057

0.061

0.0580.083

0.059

0.057

16

17

IB

IS

20

21

22

23

24

25

26

27

28

29

30

31

0.0610.052

0.057

0.060

0.058

0.052

0.055

0.C550.052

0.057

0.0570.063

-

0.054

0.0520.048

0.0570.C52

0.055

0.060

0.C58

0.062

0.057

0.0550.053

0.066

0.0570.059

mm

0.057

0.062

0.048

s s a s s

- 364 -

T i B L S - V

ANiLTSIS OF A FBf ORB SiMPLBS FflQM SINGHBHUM iRBA B I

TWO PIFFSajgHT MPHODS.

S a a p 1 e *°3°8

By ColorimetricTBP Attraction I mtf

proposed Fluorimetricmethod

Narva - A

Turaadih - A• - B

» - C

• - D

0.0470

0.0590.050

0.045

0.042

0.0475

0.0560.050

0.045

0.042

It ia clear from the above result a that the values obtained by the twodifferent aethode are aore or laoa same confirming that fliiorlaetricmethod for uranim eatiaation using alajdne-336 In benaene la a goousubstitute of spectrophotoaetric method ualng IBP extraction for theuranivai ores of singhbhtsi belt .

(6) Effect of Interferences in the letlmatJon of Oranim FluoriaatricaUy

Using Alaaine-556 aa an tebractant.

The interferences of Th, Ce, H0s" and Cl and their elimination Inthe fluorimetric eetlaation of uranim ualng 41amlna-536 in benseneas an extractant ware atodied and the results are plotted la Fig.(2)and Fig.(5). Frm Fig. (2) i t can be aeon that Ca lnterfers In uranlwiestimation in amlne ayatem of extraction but upto 1.0 g/1 of Ca can

+2be eliminated keeping the Fa concentration to 2.0 g/1 In the aquous

phase (adding freshly prepared FeflO^ solution).

- 365 -

Nitrate and Th also interfere and 1.0 g/1 of Th can be eliminated by

keeping 2.5 g/1 Chloride in aquous phase.

Nitrate was eliminated by fining vith sulphuric acid.

Prom Fig. (3) i t can be seen that Chloride does not interfere upto 2.5g/1. If the concentration goes above i t interferes very heavily. IfChloride i s present, i t should either be brought dovn to 2.50 g/1 levelin the solution or may be eliminated by fuming.

It was also observed that Mo, V, Co, Mn, Cu do not interfere upto 1*0

g/1 and Pe can be tolerated upto 5.0 g/1 .

COHCLUSIOH

Th* modified f l u o r l a e t r i o method i s most su i table for plant leach l iquors

as w e l l a s for s o l i d o r e s and t a i l i n g s samples. The preo is lon and accuracy

of the method are q u i t e sa t i s fac tory* By us ing t h i s method there w i l l be

saving o f chemicals I l k * Aluninlum n i t r a t e , Anraonlum Ni trate and Ferr ic

n i t r a t e . The es t imat ion t ime of s o l i d samples involving TBP extract ion

w i l l b * reduced, Tha consumption of solvent w i l l be very s n a i l because

t h e same solvent af t*r washing can b * reused for severa l cyc les*

The authors wish t o thank Chairman ft Mur-m ng Director, Uranium Corporation

of India limited for h i s keen interest in the Besearch ac t iv i t i e s of CRfcD

Department,

1 . Grlnaldi F.S. t- " Collect«d paper on method of Analysis for Uraniumand Thorium ". Ooo Survey Bull 1006 (1954) U.S. Govt. PrintingOffioe, Vaahington.

2 . Centanni P.A., Boss A.M. and Oesesa M.A. " Pluorimetrio Determinationof Uranium «. Anal. Chen. 28, 1651 - 1657 (1956).

- 366 -

U 3 O 8 EXTRACTION USfNGONE PERCENT ALAMINE 336 IN BENZENE

4000

o

5 3000

Ml

at

o 2000-

JO 20 J0_ .40 5Q

Ofi IN S O L U T f O N , ^

" Fl<3. 1

- 367 -

INTERFERENCE OF Th, Ce AND NO3 INAMINE EXTRACTION OF URANIUM AND

THEIR ELIMINATION

O Tb INTERFERENCE

A Ce MTgRFEKRENCE

ID NO3 INTERFERENCE

J7 ELIMINATION OF Th INTERFERENCE

© ELIMINATION OF Ce INTERFERENCE

01 02 0-3 0*4 09 0*6 Of 0*1 0*9 1*0CONCENTRATION,

no. z

INTERFERENCE OF Cf IN AMINE EXTRACTIONOF URANIUM

|100

K 90oc 80Jx 70

§50!£ 40

O.

30

2Q

f 00

O O 0

ro idoCHLORIDE ION CONCENTRATION

FIG.

- 369 -

DETERMINATION CP URANIUM IN SEA WATER BT

ABSORPTIVE DIFFERENTIAL PULSE VOLTAMETRT

R.N. Kbandekar and Badha Ragbunath

Pol lu t ion Monitoring Sect ionBhabha Atomic Research CentreTrombay, Bombay 400 065.IIMA

An adeorptiTe a tripping voltammetric procedure for d irec t

determination of trace quant i t ies of Uranium i n eeawater has been

descr ibed. Optima 1 conditions include pfl 6 . 7 , 2 x 10* M 8-hydroxy

quino l in t (Ozine) and c o l l e c t i o n potent ia l of -0.4V (Vs AgAgc l )

a t banging aercury drop e l ec trode . With control led adsorptive

accumulation f o r one Bin. a detect ion l i m i t of 2 .8 z 10 M

Uranium i s obtained. The response i s l i n e a r up to 7 i 10 M

Uranium and the r e l a t i v e standard deviat ion a t 4 z 10 MU i s 11.5J*.

The e f f e c t of p o s s i b l e interference from other metals has been

investigated.

IHTRODUCTIOM

Detexmlnation of uranium in sea waters i s of i n t e r e s t

because the element i s used f o r the production of e n e n y i n nuclear

r e a c t o r s . The contr ibut ion fvom nuclear f a c i l i t y , being sma l l ,

i s o f ten masked by the r e l a t i v e l y high v a r i a b i l i t y of uranium

concentrations in coaetal sea water. The stable oxidation state i s

uranium (vi)in oxygenated waters and i s mostly present as uranyl

ion which i s complexed by carbonate in carbonate bearing waters.

Because of the high concentration of carbonate (3 x 10 mg/l) in

sea water uranium exists predominantly (>9o£) as the trlcarbonate

uranylate anion UO_(CO,)_ ' and i t has a very high residence

time of 2.4 x 10 years w ; .

Several technique have been reported for the determination of

uranium in sea water . However these techniques are not

- 370 -

sufficiently sensitive at present for the direct determination ofuranium in sea water.

It would obviously be useful to be able to determine theconcentration of uraoiua directly in the sea water without priorcherical separation. Recently oathodic stripping voltammetrytechnique was developed for the determination of uraniun in naturalwaters* ' . This paper presents a sensitive and rapidadsorptive stripping procedure for the determination of uranium insea water using 8-hydroxy quinoline (Oxine).

Apparatus x PAR 174-4 polarographio analyser with PAR 303 bangingmercury drop electrode (HMDS) and PAR 305 electro magnetic stirrerwere used. Potentials given are with respect to Ag/AgCl,saturated XC1 reference electrode.

Reagents : i)A 3 x 10 M aqueous stock solution of U(vi) wasprepared and diluted as per the requirement. An aqueous stocksolution of 0.1 H oxine was prepared in 0.25M HCl (Analar, BOB) anddiluted with distilled water. A pH stock solution of IM PIPS(piperaiine-I-H'bis 2-etbane sulphonlc acid) mono sodium salt *nd0.5M VaOH (Arlstar, BIB) i s also prepared. An aqueous solution of0.1M EDTA was prepared from i t s sodium salt and was adjusted topH~7 by laOH.

Sea water samples were colleoted from coastal places in India.On collection in precleaned polythene bottles, they were acidifiedwith HCl so that pB of water i s nearly 2. Before measurement seawater samples were filtered through 0.45 .urn membrane f i l ters.

Procedure t A 10ml of sample aliquot was taken into thevoltammetric cell and 0.2ml of 0.001M oxine solution and 0.1ml ofPIPES pH buffer were added. After deaeration of the solution for8 min. by purging highly purified nitrogen, a fresh mercury drop wasextruded. The Sjtlrrer was started and electrolysis was done for

- 371 -

1 min. a t —0.4V. The solution was allowed to become quiescent and

the cathodic scan was carried out in d i f fe ren t ia l pulse mode with

scan rate of 5 mVs~ and the sens i t iv i ty of 500 nA(full s c a l e ) . The

measurements were repeated a f t e r three standard addition of

uranium to evaluate the concentration of uranium in the sample.

Interferences from Pb, Fe and Cd were not observed.

RESULTS AND DISCUSSION

Optimum conditions were obtained by varying the ozine

concentration, pfl of the e l e c t r o l y t e , adsorption potential and

adsorption time, scan ra te and biological buffers v i s HEPES and

PIPES.

Biological Bufferes j Effect of biological buffers on the peak

current of uranium was studied. The voltamnetrio ce l l containing

10ml of sea water, 0.1*1 of 111 HEPES or PIPES buffer solution,

0.2*1 of 1 z 10 M ozine solution was spiked with varying

concentrations of uranium (3 x 10 M to 2.1 x 10 M). After

adjusting the pH to 6 .7 , uranium peak current was recorded keeping

the adsorption tine of 1 s i n . I t was observed that better

linearity for peak current Vs concentration was obtained when

PIPES was used therefore for further experiments PIPES was

used.

Oxine concentration i The increase in peak current during

preconcentration step i s due to adsorption of uranyl oxine complex

onto the HMDE. Therefor* the peak height increase was obtained

with the Increase in oxine concentration. . The decrease in

peak current was observed a t oxine concentration of 2.5 x 10~ M.

Por analytical purpose, therefore, the axlne concentration of :

2 x 10 M was used. I t was observed that pH between 6.5 and 7 i s

quits adequate tor the above oxine concentration. The peak current

increases with the increase in pH, however i t diminishes rapidly

above pfl 7 .

- 372 -

The effect scan rate on peak current was studied (at pH 6.7)

by varying scan rate from 1 to 20 mVs" . The peak current—1remained almost constant for seen rates, 1 to 5 mYs and then

decreased with the increase in scan rate. For a l l estimation work

therefore, a scan rate of 5nVs~ was used.

Effect of changing adsorption potential and adsorption time :

In the presence of 2 x 10 M oxine and of pH 6.7 the peak

potential for reduction of uranium i s -0.68 V. The adsorptive

stripping peak height reduced considerably when the adsorption

potentials more negative than -0.4V were applied and no peak was

obtained when the adsorptive stripping scan was proceded by

adsorption at a potential negative to the uranium reduction peak.

This indicates that under these conditions U(7) does not fora

complexes having adsorptive properties.

I t appeared that peak current increased with increase in

adsorption t'me. However the increase was not linear after

1.5 to 4 minutes (at uranium concentration of 1 x 10 M). This

may be due to saturation of surface of the mercury drop.

Limit of Detection and Sensitivity t Uranium was determined in sea

water by using the adsorptive voltammetry procedure given above.

The calibration curve i s l inear up to 70 nHU (at peak current

1?0 n&) for 1 «in adsorption in stirred condition. The linear

range could be extended by using shorter adsorption time.

However, in this case the sensit ivi ty dropped considerably. The

sensit ivity obtained was 1.71 nA/nJW. The standard deviation of

the measurement of 4 nMU in synthetic sea water was 11.5jt (n>7).

The limit of detection as calculated from 3 standard deviation was

0.28 nM of uranium. This could be improved by increasing

adsorption time.

Interferences : The reduction peak potentials of adsorbed oxine

complexes of Cu, Pb, Cd and Zn are -0.47, -0.59, -0*65 and -1.02T

r e s p e c t i v e l y . These peak p o t e n t i a l s are wel l separated from that

of uranium except for Fb and Cd. However i t was observed that

they do not i n t e r f e r e with es t imat ion of uranium due to t h e i r

lower s e n s i t i v i t i e s and the concentrations in sea water. Both Pb- 4and Cd can be masked completely by add i t ion of 10 M EOTA to the

sample where as Uranium peak i s not a f fec ted

Several sea water samples c o l l e c t e d from d i f f erent l oca t ion

mostly around Bombay and few other p laces i n India during

Jan.1966 to Uarch 1989 were analysed f o r uranium using the above

standardized procedure. The uranium concentration var ies from

0.95 to 3 .95 juj l"' with the average value of 2 .36 • 0.97 )igl~1

(Table 1) the average value i s ID c l o s e agreenent with the values

obtained by o ther workers ' .

REFERENCES

1. H. Ogata, N. Inoue and H. Kalibana, (1971) , Nippon Genshirycku

Gakkaishi, 13, 560-564

2 . K. S a i t o and T.Miy&uchi, (1982) . J . Nucl . S c i . Technol, 19,

145-148.

3 . T.L. Xu, K.G. KnMis and C.G. H»thieu i (1977) Deep. Sea Res.

24, 1005-1027.

4. J . Bolzbecber and D.E. Ryan : (i960) Anal. Chin. Acta,

119, 405-403.

5. T.V. Florence and T. Parrer : (1963) Anal. Chen. 35, 1613-1616.

6 . A.M. Bond, V.S. Biskupaky and D.A. Wark t (1974) Anal. Chen.46, 1551-1556.

7 . *.C. Li, D.M. Victor and C.L. Chakrabarti, (i960) Anal. Cbm.

52, 520-523. *

8 . C.W.C. Milner, J.D. Wilson, CJl. Barnett and A.A. Snalea t(1961)

J . Electro anal. Chem., 2 , 25-38.

- 374 -

9. CM. i . Van den Berg and Z.Q. Huang i (1984) Anal. Chin. Acta

164, 209-222.

10. CM. C. Van den Berg ; (1986) Scic total Environ. 49, 89-99.

11. CM. G. Van den Berg and U. Nimmo t (1967) Anal. Chem.,

59, 924-928.

12. T.t». Sarw and T.M. Kriehnaaoorty (1968) Curr. S c i . ,

7J, 422-424.

13. C. Sreekuaaran, J.R. Naidu, S.S. Gogate, M.R. Hao, G.E. Doshi,

V.N. Sbastry, S.M. Shah, C.K. Unni and R. Viswanathan >

(1968) J . War. Bio. Asooc. India 10, 152-157.

14. B.U. Kotharl and K.C. P i l la i 1 Beport BARC/I-973, DAE India

(1979).

- 375 -

S.No.

1 .

2 .

3 .

4 .

5.

6 .

7 .

8 .

9.

10.

11 .

12.

13.

14.

15.

TABLE I

URANIUM IN SEA WATER

Place of Col lect ion

Kanyakumari, Tamil Nadu

Kalpakao, Tamil Nadu

Haei A l l , Bombay

Apollo P i e r , Bombay

Cirus, Bonbay

Thane Creak, Bombay

Tarapur Atonic Power StationCirus , Bombay

Gateway of India ( I ) Bombay

Gateway of India ( I I ) , Bombay

Band Stand, Bandra ( I )

Band Stand, Bandia ( I I ) , Bombcy

Bandz* Bombay

Versora (I) Bombay

Veraora ( I I ) , Bombay

Cone.of Uxaniumug l" water

3.571.903.813.952.14

0i95

1.07

1.67

2.66

1.90

1.07

2.362.382.62

3.09

- 376 -

DIFFICULTIES IN PREPARING A STANDARD SAMPLE OF

URANIUM METAL HAVING TRACES OF NITROGEN

R.S.D.TOTEJA, B.L.JANGIDA, M. SUNDARESANAnalytical Chemistry Division

8.A.R.C., Trombay,Bombay - 400 085

Normally in the analysis of uranium, the nitrides are

hydrolysed to give NH, and that for standardisation purposes

to approximate the closest condition of analysis of ammonia,

NH4CI is added to the sample arJ *he recovery is tested. An

appropriate method would be to have a standard sample of

uranium metal with a known amount of nitrogen to be used as

reference sample. The present work describes the efforts

made in our laboratory for the preparation of such a refer-

ence sample. Known micro-amounts of nitrogen were allowed

to react with fixed amounts of uranium metal. Since the

reaction is generally superficial, the product was homo-

genised by melting in an induction furnace. Different experi-

ments to get standards of nitrogen varying from 40 to 100 ppm

were conducted. But all our efforts met with no success to

get the desired standards. Density differences of uranium

nitride and uranium metal made the process of homogenisation

very difficult.

INtRODUCTION

The mechanical properties of uranium metal are known

to be affected by the presence of nitroqen. Furthermore,it

may get released at high operating temperatures of a reactor

tluioby causing rupture of cr.e aluminium cladding tubes due

- 577 -

to the pressure build up. Therefore„ the nitrogen content

of the uranium metal ingot is routinely monitored before

fabrication of the fuel rods. The uranium metal turnings from

several parts of the ingot are sent to the Analytical Chemistry

Division for nitrogen analysis. The samples are analysed

by micro-Kjeldahl's method . The tolerance limit for

nitrogen is around 100 ppm. It was thus essential to have

a reliable method for the analysis of nitrogen in uranium

metal at trace level and hence the necessity for having such

a standard sample arose.

( 2)Now the requirements of a primary standard are :

reasonable ease of preparation and accurate reproducibility;

purity determinable with sufficient accuracy; and stability

of the purified material under ordinary conditions of labo-

ratory. So far NH.C1 added to the solution is being used

as the standard for nitrogen determination in uranium

metal. However, the use of a standard uranium material is

advantageous because in the micro-Kjeldahl's method of

analysis the ammonia distillation and spectrophotometric

measurement in the standardisation are the same as in the

actual sample analysis and the weighing error in a standard-

isation is decreased because of the high equivalent weight

of uranium.

A standard uranium material having traces of nitrogen

is not available commercially. It was thought therefore to

prepare a standard uranium metal sample having traces of

nitrogen. This paper describes a method of such a prepara-

tion and discusses its assessment.

- 378 -

EXPERIMENTAL

a) Nitriding

Uranium metal pellets were obtained from Atomic

Fuels Division, BARC. A low pressure set up as shown in

Fig. 1 was used for nitriding uranium. About 70 g. of uranium

metal was placed in a vertical silica tube A (20 cm long and

5 cm dia). The system was evacuated and pure IOLAR nitrogen

gas was introduced slowly. An oil manometer B was used to

monitor the gas pressure in the precalibrated volume (150 ml)

of the system. The initial nitrogen pressure was kept between

70 to 150 mm of the oil manometer depending upon the desired

quantity of nitrogen gas. The silica tube was heated slowly

at the bottom in a small furnace C and the temperature was

raised to 773 K. The heating of the metal was continued at

this temperature for about 30 minutes and the final nitrogen

pressure was noted. The difference of the initial and final

pressures was used for calculating the a^cunt of nitrogen taken

up by the uranium metal. The sample was cooled and taken out.

This proceudre of nitriding was followed for four more uranium

metal samples.

b) Howogenisation

An induction furnace was used to melt the above

nitrided uranium metal sample. The sample was taken in a

graphite crucible which was kept in a glass tube closed at

one end and vacuum tight glass stopcock at the other end.

he tube was evacuated and heated to 1500 K for 10 minutes

in the induction furnace to homogenise the nitrogen content

of the sample by melting and then self-stirring. The sample

- 379 -

was cooled and analysed for its nitrogen content.

c) Analysis of Nitrogen

The combined nitrogen in uranium metal before and

after nitriding was estimated by the conventional micro-

Kjeldahl's steam distillation method . This method can

be successfully applied for determination of combined

nitrogen from 10 to 150 ppm with a variation of ^ 125S (2«") .

RESULTS AND DISCUSSION

The results are shown in Table 1. The second column

shows the nitrogen content in ppm in the uranium metal

received as such. The average value of 51 ppm was obtained

after five determinations in each case. The third column

shows the nitrogen in ppm added by nitriding. It varied

from 29 to 82 ppm. Thus the expected nitrogen content as

shown in the column four varied from 80 to 132.ppm. The

last column shows the recovery of the nitrogen in ppm. It

may be observed from the table that the recovery ia far less

than the expected values. It does not follow any particular

trend i.e. there is no direct relation between the degree

of nitriding and recovery of nitrogen.

This shows that this method is not suitable for

preparing the standard uranium material for nitrogen. The

reason for its failure may be due to two things. Firstly,

there may not have been a uniform distribution of the uranium

nitride throughout the moss of uranium since the density of

the former in lesu than lhi» latter so it would float on the

surface of the molten uranium during the process of homoge-

nisation. Secondly all the nitrogen considered to be completely

- 380 -

consumed for making uranium nitride might not have gone For

the chemical combination. Some part of it might have been

trapped inside the uranium lattice which get released on

heating in a furnace. This may explain the loss in the

recovery.

Till such further advancement in new methods for

nitrogen analysis occurs, one has to depend on the age old

Kjeldahl's method only. And since it is not an absolute

method, a calibration is a must which is done by using

NH^Cl solution only.

REFERENCES

1. S.M.Jogdeo and K.A.Khasgiwale.Report No. AEET.Anal/22, (1963).

2. R.S.Mc Bride, J.Am.Chem.Soc. 34, 393 (1912).

Table - 1

Analysis of Uranium Metal for Nitrogen

Sample Average N added Total Recovered NNumber Initial^ ppm theoreticalR ppm

PPm ppm

1 51 29 80 12

2 51 44 95 25

3 51 58 109 44

4 51 82 132 26

5 51 76 127 19

D

< D = -•-UMP

A - CONE t SOCKET

C - URANIUM PELLET

E - OIL MANOMETR

8- SILICA REACTOR TUBE

D - SMALL FURNACE

V - STANDARD VOLUME

FIG. 1 LOW PRESSURE SET-UP

- 382 -

ESTIMATION OF MANGANESE -m TAILINGS FLANT J-FPLUEMT BY ICP-ABS.

Joydeb Ray ani V.M. Pandey

QftiHIW OOBJPRjttlOa OP IHDIA UKCTH)

JADOGUDA MINES

SINGHBHUM

BIHAR

Manganese i s estimated in the tailings effluent after neutralisationwith line. Since after neutralisation at 10.00 pH, very l i t t l eManganese i s left in the tailings effluent, a very efficient methodof estimation i s required. Foraaldojdme^ method i s currently beingfolloved for the estimation of Manganese spectrophotonetrically in thehighly basic solution, At high pH (above pH 10) Manganese content intailings effluent i s so low that i t i s beyond detection Halt of spectro-photcnetric estimation, i l so , at low pH Formaldoxima method gives highresults due to the lnterferenoes of Copper, Iron, Draniisi and VanadiuD.Therefore to analyse Manganese, 1CP-JLES has been used which i s verysensitive instrument for low concentrations. It was possible to estimateManganese upto ppb lerel with 96 - 102* accuracy In plant tailingseffluent samples In IC*-AB.

IHTRODUCTIOM

In Uranium Ore prooesalng plant , i n order t o maximise Uranlun extract ion ,

a s u i t a b l e oxldant must b e added i n t h e ore s lurry during a d d or

a lka l ine l each ing . In a d d leaching process genera l ly NaClO,, 0 or

MnOg i s used a s an ox idant . I n Jaduguda Uraniw p lant , which processes

more than 1000 tonnes o f o r e p»r day, pyro lns i t e i s being used a s an

oxidant, * i c h oxldlsts tetraralent Oranlw to hexa-valant state throughFerrous - Ferrio cyle. The barren solution containing a considerablequantity o r Manganese to the tune of 1.0 gm/1, after neutralisationto pH io to precipitate msnganess and other elements, i s sent to thetailings pond. The discharge of the tailings pond generally goes to

- 383 -

the river. During neutralisation manganese i s precipitated as Mn(OU)

or finely divided MnOg. Mn(0H)2 or Mn(0H)3, i f present, are converted

by atmospheric oxidation'^'to Mn(OH) which 1B not stable in the

aqueous solutions. In coarse of time manganese may affect the environ-

ment through the media of water and soil . Of far more consequence i s

the aquatic pollution since the toxic elanent i s transported compariti-

vely rapidly^3). Direct consumption of water for domestic purposes

and Indirect assimilation through food stuff are the common mode of

health hazard. Concentration of 0,2 to 0,4 ppm are likely to cause

complaint'1 'and in general, a limiting concentration of 0,1 to 0.5

ppa has been recommended »5»6>7'.

BCPBUMBtTiL

(A) Reagents and Chemicals

(1) Cone HC1 (inalar BDH)

(2) Cone HH03 (inalar BDH)

(3) Pure Iron Turnings

(4) Pure ILOg powder

(5) Asmonlna asta - vanadat* (inalar BDH)

(6) Pur* electrolytic Copper (BDH)

(7) Pur* Calciua Carbonat* (inalar BDH)

(8) Pur* £l*ctrolytic Manganese metal

(B) Standard Solution*

Standard solution of Manganese was prepared by dissolving 99,ftS

pur* electrolytic manganese aetal in 10 a l Cone HND5 and volune

was made up to 1000 a l by adding disti l led water. The solutions

of required concentrations were made up by diluting the stock

solution,

(C) Plant Sappiest

The staples of plant tai l ings after neutralisation with l la* war*

collected .filtered and kept in polythene bottles. A suitable

aliquot corresponding to pH value was taken and neutralised by cone

- 384 -

HC1. The required volune was made up with l£ HC1 and manganese was

estimated In ICP-AES.

D. Instrxmentatlom

To analyse Manganese In low concentration 'UBTiM' make Plaema Bnisalon

Spectrometer (ICP-71D) was used. The manganese was also estimated using

SKDUDZU-W-150-02 Spectropfaotometer.

Operating Parameter for ICP

I Coolant - 16 1/min4rgon flow I Sample gas - 1 I/mln

| i iDdll lary _ 0 1/min

Power - 1.4 KIT

Torch Height - 16 on

Hare length - 257,61 m

Integration Time - 3 Sees.

Sao pie flow rate - 3.7 ml/«in.

(E) Manganese Estimation In ICP

Toe solution of manganese was taken in two different concentrations

range and were analysed in ICP-AES at the war* length of 257*61 m .

The lower rang* varied from 0.05 ppm to 1.00 pp& and the higher

range varied from 5 pps t o 25 ppa. Maasurejient of ta i l ing* effluent

for Manganess concentration ware also taken in two steps. For analy-

sing Manganese, two standard solutions having 0.5 ppa Mn and 1 ppn

Mn concentrations were used as reference. In the case of ta i l ing*

effluent having low pH i . e . high Manganese concentration, two

standard solutions having 5 ppm Mn and 10 ppa Mn concentrations

ware used as reference.

nsamxa AND DISCUSSIONS

Standardisation of Manganese Estimation in ICP.

Solutions of known concentrations of Manganese in two different concen-

- 385 -

tr at ions range were analysed in ICP-JUSS. The results are plotted inthe calibration curve of Manganese as shown in Fig.I. It i s seenfrom the calibration curve that manganese can be estimated from 0.1ppn to 25.00 ppn. Concentration of Manganese in the tailings effluentwere estimated both spectrophotometrically and in ICP-JLSS. Resultsare compared in Table-I. It i s observed that both in the ppb and ppolevel, manganese can be estimated with 95/t to 102* accuracy in ICP-AES.

TiBLB- I

* Cone of *Manganesein IGP-ABS

ppa

SI. 1 ifave length!No.t an

pH ofTailingsEffluent

II

Cone ofManganeseSpectrophoto-netrically

I Accuracy

1

2

3

4

5

6

7

8

9

10

3 s :

257.61257.61

257.61257.61

257.61257.61

257.61

257.61

257.61

257.61

Interferences!

9.5

9.7

9.79.5

6.511.110.5

3.6

4.2

9.9

3 3 3 3

0.120

0.158

0.130

0.360

212.0

630.0

416.0

0.121-

0.161

0.1280.360

208.40.0076

0.0056

620.0

410.0

0.067

100.8

101.9

98.5

100.0

98.5

98.4

98.6

The effect of interferences which are predominant in the Jfar&aldoxlaemethod were studied in ICP-AES with known concentration of Manganese,The results are given in Table-II. It i s seen from T«ble-II thatCalcium which i s most predominant species in the neutralised «ffluentcan be toleratod upto 350 ppa and Iron, Uranium, Vanadium and Copperupto 40 ppa.

- 386 -

Vaye length} Standardm

257.61•

R

R

R

257.61•

•

R

•

257.61•

•R

R

257.61R

R

R

R

257.61R

R

257.61•ftIImw

i *I "

• S3

. nEFFHCr OF IMEStFHUBfCBS

J Interfer ing>nc3ntrstion| element

of juaganese (

Ppl

1.00R

H

R

•

1.00fj

R

•

•

1.00R

R

R

R

1.00R

•

R

R

1.00•R

1.00N

R

R

*

•3 a * a a

.1

•

R

R

H

0

•

•

R

R

VR

•

R

R

OuH

R

R

R

CaR

R

*R

R

H

R

R

|

ii

Cone of Inter-fering element B

PfB

10

20

30

40

50

10

20

30

40

50

10

20

30

40

50

10

20

30

40

50

50

100

150200250300350400450

9

II

Measured coneof Ma

1.010.990.38

1.021.04

1.01

1.001.011.01

1.06

0.98

0.990.99

1*0021.004

1.01

1.001.02

0.981.061.0020.990.9651,000.990.990.960.970.96

- 387 -

CONCLUSION

Manganese vas estimated down t o t h e ppb l e v e l i n tbe neutral i sed

t a i l i n g s of Jaduguda Oranium plant i n the ICP-AES at t h e wave length

of 257 .61 m . The r e s u l t s were very much precise and accuracy varied

from 95 t o 102* .

ACKNOULEDGBiaC

The authors wish t o thank Chairman & Managine Director , Uranium

Corporation o f India Limited for h i s keen i n t e r e s t i n t h e research

a c t i v i t i e s o f C.R.&D Department.

RgRBMCBS

1 . Colorimetric Determination of Traces of Metal - E.B. Sandel. Inter

Science Pub l i ca t ions INC, New l o r k .

2 . Rankana K and SahaaaTH.G. 'Geochemistry', Dniv. o f Chicago P r e s s ,

Chicago, I l l i n o i s , USA (1950) P .640 .

3 . S tudies on Spec ie* var ia t i on of Manganese In Uranium Processing

and natural environment - M.Sc T h e s i s (Chemistry) Shree

S. Venkataranan, Boabay d i v e r s i t y 1981.

4 . Koboe R.A. Cholak, J and I*rg»nt S.J, Jour. A.U.V.A 36/1944.645,

quoted In C.A. 36 (1944) 3763 .

5 . B a y l l s , J.R. Jour. A.U.U.A 32 (1940) 1753 quoted In C.A. 36 (1941)5 4 5 .

6 . awards G.P. Jour. M.E.W.W.A.6V1947.260

7 . Neumann, R. , Z. Gesund n e l t s t e c h . S t ^ t e h y g . 25/1933 163 , quotedi n C.A. 28 (1934) 549.

- 58R -

- 389 -

VOLTAMMETRIC STUDIES OF URANIUM(VI) REDUCTION

G.A.Inamdar. and R.G.Dhaneshwar.*Fuel Reprocessing Division, Bhabha Atomic Research Centre,Trombay, Bombay - 400 085.

* Analytical Chemistry Division, BARC, Trombay, Bombay.

ABSTRACT

Uranyl reduction at mercury electrode is studied in detailunder different experimental conditions. However not sufficientdata is available for the voltammetric uranyl studies atdifferent metal wire and metal amalgam wire electrodes. Uranylreduction, therefore, was tried at gold wire, as well as, goldamalgam wire indicator electrodes, employing the three electrodesystem where platinum wire electrode is auxiliary andmolybdenum wire is reference electrode. The study was carried outin acidic, neutral and alkaline media, and in buffered andunbuffered solutions as well. In acidic medium at gold indicatorelectrode only a single curve was obtained in different acids andbuffers. The highest current of 83 jiamp is obtained in 0.1 Mhydrochloric acid for 1.0 mM uranyl concentration. Nodisproportion of U ( W couli be detected at gold indicatorelectrode. However in contrast to acidic media, two peaks wereobtained in 0.1 M each of potassium nitrate and potassiumchloride medium, current concentration linearity being obtainedfor both peaks. The highest currents were obtained in 0.01 Mpotassium chloride, being 75 and 335 jiamp respectively for 1.0 mMuranium concentration.

Similar results were obtained at S^ld amalgamelectrode though the current heights'obtained in acidic media atthe amalgam electrode are considerably smaller than those at goldelectrode. In neutral media, the results obtained at goldelectrode are comparable to the results obtained at amalgamelectrode. Thus amalgam electrode does not behave as mercuryelectrode. Comparision of the results obtained for these threeelectrodes is discussed.

- 390 -

VOLTAMMETRIC STUDIES QF U(VI) REDUCTION

INTRODUCTION

The polarographic study of uranyl ion is extensively1

reported at dropping mercury electrode. The gist of the study is:

in weakly acidic medium or neutral medium, uranyl ion undergoes

stepwise reduction giving rise to three waves. In moderately or

highly acidic medium two waves at around -0.18 and -0.9 V were

reported. There is hardly any mention of the study of uranyl

reduction at metal or metal amalgam electrodes. Uranyl reduction2

was studied at hanging mercury drop electrode, carbon or glassy3.4 5

carbon electrodes, as *ell as aluminium and platinum electrodes.

There is however no reference of uranyl reduction at gold or any

amalgam electrode. This study was therefore undertaken at noble

metal wire electrode and its amalgam electrode in order to

observe the differences if any. in the reduction pattern as

compared to the one obtained at dropping mercury electrode.

EXPERIMENTAL

The study was carried out on Electroscan-30

manufactured by Beckman Inc. "USA, employing a three electrode

system. The instrument can be operated both In the potentiostatic

and galvanostatic modes. The polarographic cell consisted of a

100 ml Pyrex glass beaker fitted with a rubber bung having five

holes for insertion of three electrodes, a nitrogen bubbling tube

and for escape of nitrogen. The electrode system consisted of

gold wire or gold amalgam wire indicator electrode(1.0 cm long,

19 SWG), platinum wire as auxiliary electrode and molybdenum wire6

as a reference electrode . Gold wir* electrode was cleaned by

- 391 -

cathodisation in 1 : 4 sulphuric acid for ten minutes. Molybdenum6

electrode is extensively used as a reference electrode.

Molybdenum wire was cleaned by rubbing with zero number emery

paper till it became shining. Gold amalgam wire electrode was

prepared by first cleaning the gold wire electrode as stated

above, then drying it and then dipping in double distilled

mercury for two minutes. "It was then thoroughly washed repeatedly

with distilled water and was then kept in distilled water for

twenty four hours for equilibration.

Stock solution of 1.0 M uranyl nitrate is prepared by

dissolving the requisite amount of Uj °a in 1:1 nitric acid. The

solution was standardised by Davies-Gray method. All the other

reagents and the acids used were of AnalaR grade or E.Merck

G.R.grade purity.

Extra pure Zolar-2 nitrogen gas supplied by Indian

Oxygen, Bombay was bubbled through the polarographic solution in

cell for ten to fifteen minutes and afterwards during the

duration of the experiment the nitrogen gas cover was maintained

above the cell solution. All the experiments were carried out at

25 +/- 0.1 degree centigrade.

RESULTS AND DISCUSSION

Gold Indicator Electrode

A) Acidic Medium :

Uranyl reduction was tried in acidic, neutral

and alkaline media. In acidic medium 0.1M acetic acid as well as

acetic acid of pH 3,4 and 6 were tried. In all these supporting

- 392 -

electrolytes, only a single peak was obtained, peak potential

being around -0.32 V (Table I). The peak currents were however

found to increase with increasing pH ; being 36 jiamp for 1.0 JBM

uranyl ion concentration in 0.1M acetic acid and 67 jjamp in

acetic acid of pH 6. Current concentration linearity was

generally observed in all these cases. However in 0.1 M di-sodium

salt of EDTA of pH 6, a S-type curve was obtained with half wave

potential around -0.23 V and the current is also reduced to 26

jiamp for 1.0 nH concentration.

Surprisingly in 0.1M nitric acid very low

currents were obtained, 30 jiamp at -0.12 V for 1.0 mM uranyl ion.

On the other hand maximum currents were obtained in 0.1M

hydrochloric acid, being 83 .uamp at -0.13 V for the same uranyl

concentration.

B) Neutral And Alkaline Medium :

Some surprising results were obtained in

neutral media. In 0.1M acetic acid of pH 7 no current

concentration linearity could be obtained (Table II). In 1.0 M

potasaium nitrate medium two peaks at -0.2 and -0.6 V were

obtained with currents of 80 and 265 jiamp. The current

concentration linearity for both the peaks is not ratisfaetory.

When the nitrate is reduced to 0,lM a single peak at -0.85 V with

the current of 290 juamp is obtained. Here also the current

concentration linearity is not satisfactory. When the potassium

nitrate concentration is further reduced to 0.01M, then two peaks

at -0.23 and -0.9 V were recorded with currents of 55 and 320

jiamp (Pig 1). The current concentration linearity here however

- 393 -

was good. The results show that as the nitrate concentration is

reduced, the currents are increasing and the results are quite

satisfactory. However, why only in the case of 0.1M potassium

nitrate only a single peak is obtained remains inexplicable. When

the study was repeated in 0.1M potassium chloride medium two

peaks at -0.25 and -0.85 V were obtained even for this

concentration; currents being 90 and 315 jiamp respectively,

.ompared to the results obtained in potassium nitrate medium the

current for the first peak is almost double while the current for

the second peak is almost the same. The current concentration

linearity is obtained for both the peaks in chloride medium. When

potassium concentration is reduced to 0.01M, the current

concentration linearity for the first peak is lost and the

currents for the second peak are not appreciably changed. It

shows that in chloride medium, unlike that in the nitrate medium,

reducing the concentration of supporting electrolyte is not

beneficial.

It is also interesting to compare the currents

obtained in nitric acid medium and potassium nitrate medium.

Compared to currents obtained in the former, the currents

obtained in the latter are ten times greater and an additional

peak having comparatively high currents is obtained. Just by+

replacing a proton by K ,such a tremendous change is obtained.

Mo proper curves were obtained in alkaline

medium.

- 394 -

Gold Amalgam indicator Electrode

\

A) Acidic Medium:

A parallel study was undertaken at gold amalgam

indicator electrode. In acidic medium, in this case also, a

single peak was obtained in acetic acid of pH 1,3,4 and 6. As

earlier in the case of gold electrode, S-type curve was obtained

in nitric acid. The deviation in the results is obtained in the

case of half wave potentials and currents. At amalgam electrode,

the half wave potential is shifted to more negative side, being

in the range of -0.4 to -0.96 V (Table III) as compared to the

voltage range of -0.12 to -0.34 V obtained at gold indicator

electrode. The currents obtained at amalgam electrode are

generally less than those obtained at gold indicator electrode

being in the range of 11 to 40 jiamp. Here also as the pH

increases currents are also increasing with.the exception of

acetic acid of pH 3. Unlike that of gold electrode, minimum

currents here are obtained in 0.1 M acetic acid and not in nitric

acid. The results indicate that the gold amalgam electrode

essentially behaves as gold electrode and not as a mercury

electrode, because at mercury electrode two or three curves are

obtained in acidic medium as noted in Introduction.

B) Neutral And Alkaline Medium :

The study was carried out in different

concentrations of potassium chloride and potassium nitrate

medium.. In 1.0 M potassium chloride a single peak is obtained at

-1.05 V; with a current of 335 pamp. The current is linear with

- 395 -

concentration and when the concentration of chloride is reduced

to 0.1 M two peaks are obtained at -0.29 and -1.12 V with

currents of 55 and 400 jiamp respectively, both currents varying

linearly with concentration. When the chloride concentration is

further reduced to 0.01 M the first peak at -0.29 V disappears

and a single peak at -1.18 V is obtained with a reduced

current of 300 jjamp. In the case of 0.1 or 0.01 M potassium

nitrate supporting electrolyte, a single peak is obtained around

-1.10 V with currents at around 300 jiamp. Changing the

concentration of nitrate has only marginal effect on peak

potential or current.

Summary and Conclusion

1) Only one uranyl reduction curve is obtained at gold

and gold amalgam electrode in acidic medium. In acidic medium as

pH increases current is also increasing. Minimum current is

obtained in 0.1 M nitric acid for gold and 0.1 M acetic acid for

gold amalgam electrode. Gold amalgam electrode behaves as gold

electrode and not as mercury electrode, only difference being the

more negative potentials for the amalgam electrode. The currents

are also a little lees than gold electrode.

2) Two peaks were obtained at both gold and gold

amalgam electrodes in neutral medium i.e., pot issium nitrate or

chloride ; the second peak being ten times greater than the one

obtained in acidic medium. However, in acetic a»~Ld. medium of 7.0

pH only one peak was obtained. The extent of the current as well

as current-concentration linearity is dependent upon the

concentration of potassium nitrate and to a lesser extent

- 396 -

potassium chloride. At lower potassium nitrate concentration, the

currents for both the peaks increase and the current-

concentration linearity becomes very satisfactory. The most

noteworthy fact is that nitrate ion catalyses the uranyl+ +

reduction when it is associated with K and not with H . So is

the case for the Cl .

ACKNOWLEDGEMENTS

The authors gratefully acknowledge the constant

encouragement given by Shri. M.K.Rao., Head, F.R.D., B.A.R.C. and

Dr. R.K.Dhumwad., Head Laboratory Section, F.R.D., B.A.R.C.

during the course of this work.

REFERENCES

1. I.M.Kolthoff and J.J.Lingane,"Polarography",Vol 2ndInterscience Publishers, New York,1952.pp 462 et seq

2. J.Ferreria, S.Batstachaves, M.Fatima, A.AbraoChem. Abstr. Vol 108 (1988) 153849

3. K.Izutsu, T.Nakamura, T.AndoAnal. Chim. Acta,152 (1983) 285-8

4. K.H.Lubert, M.Schnurrbuschibid,186 (1986) 57-69

5. C.A.Harte, B.P.SanchezChem. Abatr. Vol 94 (1981) 202112q

6. V.T.Athavale, (Mrs) M.R.Dhanesfiwar, R.G.Dhaneshwar;J. Electroanal.' Chem ; 14 (1967) 31-35

7. W.Davies, W.Gray,Talanta, 11 (1964), 1203

- 597 -

TABLE 1

URAWYL REDUCTION _{ ACIDIC MEDIA hi GOLD ELECTRODE

Apparatus : Electroscan-30 Electrode System : Au/Pt/MoVoltage Scan Rate : 40 mV/sec

Sr.No.

1

2

3

4

5

6

7

8.

9

10

11 !

12 I

13 !

14 i

1 2+!UO Cone.I 3! mM

1 0.5

! 1.0

! 0.5

1.0

0.5

1.0

0.5

1.0

0.5

1.0

0.5 1

1.0 I

0.5 !

1.0 !,

1 Supporting Electrolyte

t M

1 Acetic acid,

i *

1 Acetic acid,

Acetic acid,

M

Acetic acid.

n

Disod. EDTA.

tt

Nitric acid,

n

Hydrochloric

*•

0.1

H

0.1,

H

0.1,

n

0.1.

H

0.01

0.1

H

acid,

pH 3

pH 4

11

pH 6

pH 6

II

0.1 I

H 1

I11t

! -0

1 -0.

1 -0.

1 -0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

Ep

V

34

37

29

32

32

31

_—__— _36_________26—— —22

24

15

12

18 !

13 1

1 iP

1 jiamp

1 20.5

36.0

24.5

47.5

25.5

47.0

31.5

— -—67.0

12.0

26.0

15.0

30.0

39.0

83.0 !

I Remarks

1 Peak !

1 " I

Peak !

it 1

Peak !

I

Peak !

>< 11

S-type 1

!

S-type I

II l

Peak !

M 1

- 398 -

URANYL REDUCTION J

Apparatus"oltage Scan Rate

TABLE II

NEUTRAL AU£ ALKALINE MEDIA AJ£ GOLD ELECTRODE

Electroscan-3040 mV/sec

Electrode System : Au/Pt/Mo

Sr.No.

1

2

3

4

5

6

7

8

9

10 !

11 !

12 »

I 2+!UO Cone.! 2t mM

1 0

I 1

0

1

0

1

0

1

0

1.

0.

1.

.5

.0

.5

.0

5

0

5

0

5

0 !

5 !

0

1 Supporting Electrolyte

1 M

1 Acetic acid.

Pot.

Pot.

Pot.

Pot.

Pot.

H

Nitrate,

M

Nitrate,

Nitrate,

Chloride

Chloride

H

0.1, pH 7

H It

1.0

ft

0.1

0.01

II

0.1

II

0.01

II 1

1

1 -0.

1 -0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-0.

Ep

V

39

38

20,

22,

84

85

20,

23,

20,

30,

22,

31,

-0.

-0.

-0.

-0.

-0.

-0.

-0.

-1.

71

62

85

90

80

92

90

04 '1

1

34

80

28

55

40

90

46

75

ip

jiamp

26.5

34.0

.0,170.

.0,265.

176.0

290.0

.0,156.

.0,320.

.0,154.

.0,315.

.0,154.

.0,335.

0

0

0

0

0

0

0

0 I— !

! Remarks

Peak

Both Peaks

Peak

n

Both peaks

ii

Both peaks

Both peaks

For alkaline solution (Disod. EDTA, 0.01 M, pH 8 ) no proper graphs.

- 399 -

URANYL REDUCTION

Apparatus : Electroscan-30Voltage Scan Rate : 40 mV/sec

TABLE III

ACIDIC MEDIA A£ GOLD AMALGAM ELECTRODE

Electrode System : Au(Hg)/Pt/Mo

Sr.No.

1

2

3

4

5

6

7

8

9

10

2+UO Cone.

2mM

0.5

1.0

0.5

1.0

0.5

1.0

0.5

1.0

0.5

1.0

Supporting

M

Acetic acid

Acetic acid

it

Acetic acid

H

Acetic acid

Nitric acid

Electrolyte

. 0.1

.. 0.1.

It

, 0.1.

n

, 0.1.

, 0.1

II

pH 3

ft

pH 4

II

pH 6

•I

Ep

V

-0.38

-0.37

-0.90

-0.96

-0.40

-0.40

-0.42

-0.35

-0.72

-0.70

ip

jiamp

11.2

19.6

6.4

13.0

13.0

25.5

16.0

42.5

15.5

31.5

Remarks

Peak

»

Peak

it

Peak

it

Peak

S-type

- 400 -

TABLE IV

URANYL REDUCTION U£ NEUTRAL AND ALKALINE MEDIA AT. GOLD AMALGAM ELECTRODE

Apparatus : Electroscan-30Voltage Scan Rate : 40 mV/sec

Electrode System : Au(Hg)/Pt/Mo

Sr.No.

2

"3

4

- 5

6

7

8

9

10

UO2+Cone.

2 1mM

0.5

1.0

0.5

1.0

0.5

1.0

0.5

1.0

0.5

1.0

Supporting Electrolyte

M

Pot.

Pot.

Pot.

Pot.

Pot.

Chloride, 1.0

Chloride. 0.1

n i»

Chloride, 0.G1

n

Nitrate, 0.1

H ft

Nitrate, 0.01

Ep

V

-1.07

-1 .05

-0.26,-1.08

-0.29,-1.12

-1.10

-1.18

-1.00

-1.04

-1.13

-1.19

ip

jjamp

174.0

335.0

24.0,182.0

55.0,400.0

142.0

300.0

158.0

295.0

144.0

315.0

Remarks

Peak

Both peaks

Peak

Peak ^

tt

Peak

Peak1

For alkaline solution ( Disod. EDTA. 0.01 M, pH 8 ) no proper graphs.

- 40'. -

7-

6 -

oI

J -z 3

or

3

Apparatus; Electro scan - 30Electrode System ; Au/pt/^4o

Voltage Scan Rate : 4Om t f / s f 5Current S@nsHiv<tyCurve A - 1.0 mMCurve B = 0-5mM

UOa-l-~ SGfBflmp

U0

4- 0-2 0.0

URANYL

- 0-2- 0 . 6 - 0 . 8 - ) ' V - i • * - '••«

VOLTAGERg - 1

REDUCTION IN 0,01 M POTASS.UM NITRATE AT GOLD WIRE ELE C T RODE

- 402 -

Session III B

DISCUSSIONS

Paper No. 2

S. GANAPATHY IYER s In normal fluorimetrio method the calibrationgraph is constructed for the rarge 0,01 - 1ug U. In the projectedcalibration curve the concentration range i s given as 10-50ug.I fee l there i s acme oversight in this*

A.B. CHAKRABORTY t The range shown i s what if ^resent in 1Cml ofthe solvent phase. Since 0.1ml of this extract i s taken foractual fli'orimetric analysis, the range for calibration mayplease be read as 0.1ug to 0.5ug U,Og. Since most ' the samplesanalysed by fluorimetry in our laboratory f a l l in itua range ofuranium concentration, the calibration graph given is also for'this range. The amlne concentration (1£ antne) has been foundto be suitable tor this range*

Paper Wo, 5

B.L. JANGIDA t What la the significance of determining of Mnin uranium tailIng3?

JOYDEB RAY t During neutralisation, manganese Is precipitatedas Mn(0H)4 or finely divided Mn02Mn(0H)2 or Kn(0H)?. If presentaa Mn0g.Mn(0H)2 or Mn(OH), it gets converted to Mn(0H)4 byatmospheric oxidation which la not stable in aqueous solution*In course of time Mn may affect the environment through themedia of water for domestic purposes, and indirect assimilationthrough food stuff are the common mode of health hazard whichaffects mainly the central nervous system. Therefore astatutory level 0.5 ppm in the neutralised tailings planteffluent has been recommended*