Proceedings of the International Symposium on URANIUM TECHNOLOGY

VOLUME I

BHABHA ATOMIC RESEARCH CENTRE, TROMBAY, BOMBAY, 400 085

DECEMBER 13-IS, 1989

Organised by ENGINEERING SCIENCES COMMITTEE BOARD Of RESEARCH IN NUCLEAR SCIENCES DEPARTMENT 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 CM. 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 at BARC, Bombay during Dec. 13-15,1989 under the auspices of Engi- neering Science Committee of Board of Research in Nuclear Science, Department, of Atomic. Energy.

In the context of expanding nuclear power programme in Indiai the need for production of large quantity of uranium fuel from indigeneous resources hardly needs any elaboration. After three decades of experience in this area, it is thus appropriate to pool together the expertise and experience gained in the country 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 an opportunity to the uranium technologists to interact and exchange their experience and plan future strategies.

Being the first symposium on this toppic the response has been excellent as evidenced by the extensive participation and contribution of papers covering almost all aspects of uranium technology. Altogether 69 papers, including invited lectures have been presented and the proceeding* have been brought out in two volumes.

It is our hope that this technical coverage of the pro- ceedings and panel discussion would serve as a valuable reference material for uranium technologists in the coming years; We wish to record our sincere thanks to the members of the organising and various other committees, authors of invited lectures, and contributed papers, and panel members for making it possible to bring 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 Beharl t

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. Vem, 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 A21 reserves 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 a o 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 «• MV 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. A32

TABLE - I

URANIUM RESOURCES OF MAJOR PRODUCING COUNTRIES

Data * as on 1.1.1981, cost range US $ 130/kg U Country Reasonably Assured Estimated Additional •000 t U •000 t U

Australia 317 285 Brazil 119 81 Canada 258 760 France 74.9 46.5 India 32 25 Namibia 135 53 Nigeria 160 53 South Africa 356 175 U.S.A. 605 1,095 Others 237 147

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)

Estimated

Country 1987 1988 1995 2000

Australia 3.8 3.6 8.7 9.9

Canada 12.4 12.4 14.8 17.7 Europe (Mainly 3.7 3.8 3.1 1.4 France) Naalbla 3.5 3.5 4.0 3.5 Gabon Niger 3.8 3.9 4.5 2.8 (Central Africa) South Africa 4.0 3.8 0.77 0.77 U.S.A 4.8 5.2 5.7 4.9 Others 0.7 0.6 5.6 6.4

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 0 2 5

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

3,60,000 135 Wyoaing Mineral Corp. Florida 4,50,000 160 Gardialr Incorp. M 5,00,000 170 International Minerals H 7,60,000 290 &Cheaicals Corp. •t 1,190,000 485 Earth Sciences, Inc. Alberta, Canada 145,000 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

Capacity Plant „, 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 II A

URAHIVM PROSPECTING

Chairman : SHRI S.3A3TRY Chief Geologist UCIL. STRUCTURE AS A GUIDE FOR URANIUM EXPLORATION IN THE TURAMDIH MOHULDIH AREA, SINGHBHUM DISTRICT, BIHAR

R. MOHANTY and M.B. VERMA Atomic Minerals Division 34, Khasmahal, Tatanagar - 631 002

Uranium mineralisation at Turamdih is hosted by chlorite- quartz schist±apatite and magnetite* whereas at Mohuldih, it occurs in the immediately underlying quartzite and tourmaline- bearing sericite schist. The ore horizons are in the form of a number of lodes, concordant with the schistosity of the host rocks, and separated from each other by a few tens of metres of 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 is mostly affected by the F2 fold movement. Critical information on such structural guides for mineralisation will help in pla- nning evaluation drilling programmes in the contiguous area to substantially augment the presently-known reserves of uranium* -2-

INTRODUCTION

The Singhbhum Shear Zone (SSZ) is well known for its Cu-U mineralisation. Though it 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, lies 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 . I

Rock type Stratigraphy as Known uranium referred by occurrences Dunn and Dey, 1942

N Game ti ferous Chaibasa Stage Bagjata*, ^ mica schist and Kanyaluka, guartzltes Gohala Chlorite schist Iron Ore Stage Narwapahar* + apatite and Turamdih* magnetite Garadih* Rajgaon . Sericite schist ± .do- Mohuldih* tourmaline Bangurdlh Banded magnetite quartzite S Soda granite Soda granite

* economically viable deposits

, The local structure is in no way different from the regional structure described by Hikhopedhyay (1964). As described by him* there are three folding episodes decipherable in the area* In a broad sense, the earliest deformation (Fj) is of tight to iso- clinal reclined folds with the development of a pervasive axial plane shlstoslty (*x) which, at most places, is parallel to the bedding (So). The general trend of the foliations is MM# - SSI dipping 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 affacted by later extensive mylonltisatlon that small scale Fj folds have become scarce* The down dip llneatlons which are profusely developed on S^ planes parallel the Fj fold axis and hence are

Tx lineatlons (L1). Incidentally, these lineatlons also parallel the strlatlon lineations (a-llneatlons) pertaining to the later phase of folding* However, F^ folds in the mappable seal* are sometimes preserved In the quartzite outcrops (rig.2)* -4- The second generation of folds (F ) trends ESB-WNW and are nonplunging to low plunging either towards east or west. Most of the small scale folds visible on the surface belong to this generation. The earlier schistosity S. has been affected by s this folding. A set of crenulation cleavage C 2^ Parallel to its axial plane has been developed more dominantly in the schis- tose rocks. The lineations (L_) pertaining to F~ folds occur in the form of puckers on the S^ surface. Not much variation in attitude of F~ folds is seen because the S- surfaces are

fairly consistent in their attitude and F1 hinges are very rare.

Overprinted on them are the P^ folds in tne form of broad warps with very high wavelength/amplitude ratio. The axes of these folds trend almost N-S with moderate amount of plunge towards north. No small scale manifestTTations are, however, recognisable except minor strike swings. The effect of these three generations of folds on the ore body in the subsurface are discussed in the following.

URANIUM MINERALISATION

Host rocks The uranium occurrence, at Mohuldih and Turatndih was known during late fifties (Bhola, 1965). Mineralisation at Turamdih is hosted by chlorite quartz schist ± apatite and magnetite whereas that at Mohuldih is hosted by the immediately underlying unit of sericite schist and banded magnetite quartz!te (Table-I). The chlorite schist at Mohuldih, which is in the strike continua- tion of 'Airamdih, however, contains impersistent uranium horizons.

Exploration

Exploration by evaluation drilling at Turamdih and Mohuldih has been carried out through various stages during the last three decades. 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 it is 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 its 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 lies about 2 km west of Turamdih, minera- lisation occurs on surface over 350 m strike length with sub- surface continuity of little 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. It is of interest to know how each of the three folding episodes described earlier, has affected the uranium lodes in the subsurface. While the mineralisation is confined to one particular lithic unit, it occurs in the form of a number of layers and is 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. Series Dip amount and direction

1. I and II 30° towards N30°E 2. II and III 30° towards N35°E 3. III and IV 20° towards M40°E 4* IV and V 15° towards N70^B 5. V and VI 28° towards N80°I 6. VI and VII 24° towards S70°S 7. VII and VIII 26° toward* S70°E 8. VIII and IX 30° 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

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 is 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 critically reviewing the paper*

REFERENCES

Bhola, K.L. (1965) » Radioactive deposits in India. In 'Uranium Prospecting and mining in India1, 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. Sci., V.3, p.208-219. Mukhopadhyay, D, (1984) * The Singhbhum Shear Zone and its place in the evolution of the Precambrian mobile belt of north Singhbhum. Ind. Jour. Earth Sci., 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 ZONE SHOWING- URANIUM DEPOSITS DISTT. SWGHBHUM

PIG 1 - 11 -

OEOLOOICAL MAP WITH BOREHOLE LOCATIONS , MOHULDIH .

DISTRICT - SINOHBHUM , BIHAR

M> O SO CO SCALE

; MOMUUJIH CAM*

IUCK CHLOWTt SCHIST. K0OIN0 m l m\*\ CHLMITC SCMCIU SCHIST 5CHUTOV1Y IV.-.I WWOTC 9CMST WltH TOUMMUNS LOCATION Of •OftCHOLfl I.V.M •WOO MMWTITI OUMTtnt [•».».«! SOM OWUMC

FIG 2 - 12 -

BOREHOLE LOCATION PLAN OF NANDUP-TURAMDIH AREA

x \ \ C ,J

AJL FIG 3 - 13 -

150*- b^l l f — 100 • so - 0-0 • 50 1 I 1 7 6A 6 1 t 200

|>r-;.| 3E/tICITJt 0UAJIT2 SCHIST f- 1 URANIUM Oft* prrren CKUXIITC QUARTZ SCHIST/FSLU6PATHIC SCHIST

B

Pis. **> Vertical FroJ*«tloa of Ore »t Soutk aloBg AA*. Fit. LoasituAlMl Vertical irojcctlo* of Ore. ko*j ait Turaailk South aloai !*•. (r«ftr*a«« !!•• *r»w« la - H -

FIG 5b - --,| Sariaita Quartz Sakiat TH Cklorlt* Quartz Saaist i..' I Quartzita a*a Serlaita Sakiat Uramlua Ora- K;:-»'*fl (tal«o««) witk touraallma It .25 22 2f . ISO H

FIG sa Strlka Saatlbm aloaj tka aorakolas of III Saries at Mokulalk. Fie* Sat Strlka Saatloa aloac tka korekolaa of IX.S«rita at Mokulalk. ( rafaraaaa llaa aratm lm Fie? )• FIG 6 ?2 folis _o» tke «ore of the borekoles of Mohuldlk. - 16 -

FIG; ^^.^-j Seri«ite Qttartz ScJ»i»t ^ j Chlorite Quartz S

I.1. .• J Suartzite a*4 Serivlte S«alst Uraalua Ore 1 " 'n (tal«cs€) wita tour«alia«

A Typical Dip Se«tiom aloag tke ¥oreliol«i of Kohuliik. ( reference liae Arawa i» Pit.2 ) - 17 -

L-NANDUP TUCAMOlH SOUTH TURAMOIH NORTH m 2>. 197

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»« lia^lrim lm Pl«.3 ) 6 01J

V«0 NUU otrvenn tv «o jtmi

TTHWOOI jo iaviM»

' HVHM' IWWH9WS ISM 'V3UV HKrWHON *m W01N03 TWUMUUS it1* oe

- 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

and M. THIMMAIAH

Southern Region, Regional Centre for Exploration and Research, Atomic Minerals Division, Department of Atomic Energy, BANGALORE -72

Detailed exploration of the carbonate rocks of the Vempalle Formation of the Proterozoic Qiddapah Supergroup has led to the identification of a promising stratabound uranium horizon having correlatable mineralization of good width, grade, and extent in 18 localities over a stretch of 62 km. Sub-surface exploration at two localities (Thummalapalle and Gadankipalli) has resulted in delineation of ore bodies with good grade* sizeable tonnage, and thickness down to shallow depths of about 150 m.

The exploration methodology adopted, various Integrated techniques used and guides recognized during exploration* together with results obtained are discussed. Suggestions for developing exploration programmes in similar lithostratigraphic settings elsewhere in the country are also made* - 20 -

INTRODUCTION

The middle to late Proterozoic Cuddapah basin has been a favourite ground for exploration geologists and mineral prospec- tors since as early as 1625, due to its associated diamond occure- nces as well as asbestos, barytes, and base metal mineralisation. This basin has been radiometrically surveyed since mid sixties because many favourable criteria for uranium concentration like the stratigraphlc setting consisting of middle Proterozoic psammo- pel i tic sediments and chemical precipitates, very fertile grani- toids in the vicinity, and the repeated phasas of igneous acti'ity of both basic and acidic nature, are present. During these earlier surveys, the basal Gulcheru conglomerates resting unconformably over the Archean gneisses/granites were found to be radioactive mainly 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 Atomic Minerals Division have brought to light a unique type of strata- bound U-mlneralisatlon in association with the Vempalle carbonate rock belonging to the Middle Proterozoic Papaghnl Group of the Cuddapah Supergroup* The mineralised carbonate rock is admixed, at many places, with phosphatic and siliceous material, and has been traced over a stretch of 62 km, wherein about 18 interesting zones are delineated by ground radiometric surveys (Fig.l).

After the discovery of this mineralisation, a systematic exploration methodology has been adopted, taking Into considera- tion both the field guides and genetic models. Photogeology, airborne gamma ray spectrometrlc techniques and hydrogeochemical surveys were adopted during the early stages to cover larger areas in shorter time and to delineate favourable areas for detailed follow-up investigations. Results of the hydrogeo- chemical surveys carried out In the Vempalle Carbonate rock and the overlying Upper Cuddapah sediments are discussed in a separate paper presented in this Symposium. In the anomalous zone*. -21-

detailed ground checking by radiometric methods was followed by shielded probe logging of the outcrop areas. Encouraged by the good strike extensions, width* and favourable analytical results, shallow down the hole drilling was initiated in the later stages. After the Initial success, core drilling was introduced 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 of encouraging grade and thickness, together with its geological set up are dealt with in this paper.

GEOLOGICAL SET-UP

Different aspects of the Cuddapah basin are described In the classic work of King (1872). MLth an object to provide • new outlook in understanding the evolution of the Cuddapah basin, a revised llthostratigraphlc classification has bean proposed recently by Nagaraja Rao et al., (1987) taking into consideration the stratigraphy, structure and evolution of the basin.

. This stratlgraphlc succession and the uraniferous horizons identified in the Lower Cuddapah sediments are given below*

Age Group Formation Rock types ? Gandikota cju*rtrite Cuartrite Tadapatri shale Shale Rilivendla quartzite Conglomerate U-minera- Y Ouartzite llsation Li msconformity * Wtmpalle limestone/ Stromotolltlc (U-minera- n z o-r>«K«4 shale dolomite, (llsation °xi jO %2jjwrw2jpjl 1 stonedolomita, chert, mu,d (CStratabound) s C breccia basic J silla and dykes t Gulcheru quartsita Conglomerate, Unconformity ^y^^i^t Arch- ttMmmmmn*' QCV\LXB/ \ (u-mlncra- aean »asem»nc Gneisses I lisation \ (fracture I and shear I controlled) - 22 -

The Vempalle Formation, which is by far the most important from the point of uranium mineralisation, conformably overlies the ^ulcheru quartzite, both constituting the Papaghni Gro\(p of the Lower Cuddapahs.

Gulcheru quartzite, the basal member of the Papaghni Group overlying the Archaean basement (ilg.l) with a profound uncon- formity, consists mainly of conglomerate/grit, arkose and quart- zite with shale intercalation, and has a thickness of 33 to 280 m.

The Vempalle sediments are mainly calcareous consisting of stromatolites and dolostone with intercalating quartzites, con- glomerates and chert bands. The estimated thickness is around 1800 - 2100 m (Roy, 1947) • This unit is traversed by basic dykes. Lower Cuddapah sediments have witnessed magmatic activity manifested in the form of sub-aerial basic lava flows, sill and dyke intrusions.

EXPLORATION METHODOLOGY AND RESULTS

Ground radiometry

Initial ground radiometrlc checking has revealed the presence of mineralised carbonate rocks (Vempalle Formation) recording radioactivity of the order of 3 to 10 times the back- ground count and commonly rising above 15 times intermittently along a 62 km long belt between Komantula in the west and Cuddapah in the east. Eighteen anamolous zones have been identi- fies in this belt, with individual outcrops varying from 200 a to 1.5 km in strike length and 20 to 25 m in width. The important localities* where detailed investigations are being carried out, ares Tummalapalle, Gadanklpalle, Rachakuntapalle, and Bakkanna- garipalle. It has been noticed that high order radioactivity in the carbonate rock is associated with silica-rich portions, dark bands of slltstones, chert and stromatolites. Radiometrlc assay values of about 200 grab samples from these areas show 0.01% 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 drilling

Shielded probe logging of the mineralised outcrops over the dip slopes indicates average values of the order of 0*02% u to 0.03% e 3O0 over widths of 10 to 25 m. As most of the rock exposed is along the dip slope and escarpment outcrops are lacking, shallow down the hole (DTH) drilling is carried out to know the true thickness and grade of the mineralised horizon in two promising areas - Tummalapalle and Gadankipalle - with a drilling interval of 50 m to 100 m along the strike to intercept the mineralisation at a depth of 10 m to 30 m. With this drilling, a strike length upto 1200 m each is delineated both at Tummalapalle and Gadankipalle. The grade and thickness of the mineralisation varies from 0.02% u •U308 x 1.5m to 0.050% « 308 x 4.5 nu

Core drilling

In order to study the subsurface samples with respect to mineralogy, grade and geochemical parameters, core drilling is carried out In these two areas. The pattern of borehole locations and results obtained from each of the two areas are given below*

Tummalapalle deposit

In the Tummalapalle area, the uranlferous carbonate rock is sandwitched between a lower massive limestone (with intercalatory shale bands) and upper cherty limestone (Fig.2). The mineralised carbonate rock measures upto 20 m, and is further made up of inter- calatory mudstone, with development of mudcracks. At places* ripple marks and stromatolites are very common in this carbonate rock* A thin lmpersistent layer of conglomerate is often recog- nised separating the underlying massive limestone and the uraniferous horizon. The shale unit immediately succeeding the mineralised zone Is 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 strike, with low dips of 10° - 15° towards north (Fig.2).

In the first 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 EOr 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 is 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 is 20-30 m. The basic dyke* which is so prominent at lUmmala- palle, is 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 is selected for core - 25 - drilling in 100 m x 100 m grid to know the depth persistance of mineralisation. Drilling completed so far has thus indicated correlatable ore grade mineralisation, both along the strike and dip.

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 is 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 bands of dolomite-rich carbonate and collophane-rich phosphate. The radioactive minerals - pitchblende, and coffinite - occur either within the phosphate-rich band or at the junction between this and the carbonate band. In addition* some suspected organic material in association with pyrite has been identified. The P-Oc content in the surface samples varies from 5 to 15%, and upto 35% very rarely, whereas in core sample it is 1 to 5%. p in There is good positive correlation between uranium and 2°5 core samples, whereas the sane in the surface samples is insigni- ficant.

Among the trace elements, Ni, Cu, and Mo are present in appreciable concentration as compared to the Clark's values. Of these, 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, it has been established that the uranium minera- lisation, confined to the carbonate rock of the Vempalle Formation, is str at abound. This lithic unit occupies a distinct stratigra- phic position being sandwitched between the massive limestone and the cherty limestone of the Papaghnl Group. This strati- graphic control and other associated sedimentary structures like ripple marks, mudcracks and stromatolites are important field guides for locating the uraniferous horizon in the study area.

Exploration by non-core and core drilling methods in the Tummalapalle area has resulted in delineating a cor rel a table and continuous ore zone of over 1.8 km, thereby establishing substan- tial 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 drilling of non-coring and coring methods judiciously, the evaluation has been made possible in shorter time, besides economising the drilling cost to a great extent. The correlatability of ore bands both along the strike and dip and the high degree of consistency of their grade and thickness are remarkable in the two study areas, further drilling in these two areas is in progress to establish additional reserves in the inferred category and to convert the rmamrvmrn from inferred to indicated category. The high concentration of molybdenum in the ore zone (average about 0*03%) is an additional factor to enhance the economic viability of these deposits. Another very encouraging aspect is the disequilibrium factor, generally in favour of parent uranium of order of 20-25%, which would enhance the 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. It is 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 the Vempalle carbonate rock with vast lateral extent and remarkable consistency in grade and thickness has the potentiality to contribute substantial 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 exhibit similar llthostratigraphlc and chronostratlgraphic characters as of the Cuddapah basin. It is hoped that a systematic study of these basins on the lines carried oat in the Cuddapah basin, would bring 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., RAMALINGA SWAMY G., and RAVINDRA BABU B. (1987) x Stratigraphy, structure and evolution of Cuddapah Basin. Geological Society of India, p. 33-86. ROY A.K. (1947) t Geology of the Ohone Taluk and neighbouring parts, Kurnool district. Geol. Surv. India progress Report (1945-46) (unpublished) - 30 - btOI.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:

713 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 3I1ALP W.:~::^l f'ULLIVENOLA/NAGARI QUAMTZITE n=^' MASIC SILLS/VOLCANIC FLOWS i'•.'•'• IVF.MPALLJ: DOLO5JONC/LIMLSIUNI£/

GEOLOGICAL MA? OF TUMMALAPALLE AREA CUODArAH. DISTK A.P).

, k*ii: «;jr. I CM - 35 -

TUMMALAPALLE AREA CUDDAPAH. DISTT. (A.P)

•lti.it Qt

i^^Ml*s^ve umt sroue GEOLOGICAL MAP OF GADANKI PALLE AREA CUC3APAH.DI5TT. (A.P)

3 13D

1 _"2 •: 1 ' 1 -• *' c-H 1 1 1

P I [ CMERTV LIMtSTONt , CUDDAPW OlSTT

^ 3.5 » .i :i;..|i':i^v IT «W.U 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, viz., '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. A qualitative analysis of these lineaments indicates presence of linear magnetic sources having mafic to ultramafic composition at very shallow depths. Many NW-SE faults either terminate or laterally shift these lineaments at several locations. Thus, both the structural trends, i.e. E-W and NW-SE are recognised on the aeromagnetic maps of both the blocks. Uranium anomalies from airborne spectrometric data have been plotted on the aeromagnetic maps. The distribution of these anomalies indicates that uranium mineralisation has a preferential enrichment close to the NW-SE structures, contact zones and near the intersection of E-W and NW-SE structures. It is. therefore, concluded that NW-SE structural features and contact zones may be promising targets for ground follow-up.

INTRODUCTION Aeromagneticshas a long history as a method of geophysical exploration and Is a very Important tool used In any mineral or oil exploration programme. Besides delineating structural features and lithologlcal units, it plays a significant role in oxploring and identifying potential mineral belts. Recently, Grant (1985) has given the geophysical concept of "Ore Environments" that can be recognised from airborne magnetometor surveys due to the characteristic features of magnetic mineralogy. Though, thoro may not be a direct relationship between magnetite and uranium ore environment, combination of aeromagnetic and aerial spectrometric data may identify potential areas of uranium ore concentration. The present study deals with the interpretation of aeromagnetic maps of two areas in Madhya Pradesh, bordering the Chhattlsgarh Cuddapah Supergroup. Onn area in the 'Bilaspur - 38 -

Block1 is north of the Chhattisgarh basin and the other in 'Raipur Block1 is located south of the basin (Figure 1).

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 with bauxite

Upper Cretaceous Deccan Traps Fissure lava flows

Upper Carboniferous Talcher Groupt Boulder bed. (Lower Gondwanas) conglomerates, needle shales 8 sandstones

Upper Protorozoic Raipur Group Limestones fi shales

(Chhattisgarh Chandrapur Group Sandstones Cuddapah Suporgrnup)

Lower Proterozoic Granites, dolorite dikes, schists

Archaean Grant to - gnolsses, schists, amphlbolltes M4? »I .4 4 .4 * Arttt «•»•

LEGEND

L«Mtt«« Cn.Cl9T0CCHC>

(T) DltcM Trty CCMCTACCOUS-COCCHC)

(T) limit• CrMp (UCUCCOUSI

(T) U«tr *MrfwM«i (UfPCR CARtONITCROUS- V~^ 10WC* TNIASSIC )

(T) CklM«llM«rl> C«««,iKt(UfPER PROTEROZOIC)

TCROZOIC ) (T) U»cU«lil.«

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 strike, 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 -

.;%• • ^

0 jgw/MMuj* 1 © Ml CM >M> © t~..._* O MM»«UMf< j © © e 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 field, 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 -

Fig.3 TOTAL INTENSITY AEROMAGNETIC MAP OF vW*va •••••. 4« . ^ 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 of uranium anomalies 4 seen at Akalwara and down south all along the contact zone between the Archaean and the Chhattisgarh Cuddapah.

DISCUSSION OF RESULTS AND CONCLUSIONS The 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 have been deduced from their magnetic signature that is duo either to the lateral displacement of "magnetic horizon" or its discontinuity. These faults are parallel to the major regional Mahanadi tectonic trend. Domzalskl (1966) discusses that the dikes represent moat important structural features that can be related to the major directions or fracturing. P. Gay (1972) also observed that the nornmngnotlc llnonnmnts c:nn bo correlated with major tectonic events. Thus, the E-W magnetic lineaments and dikes In Bllaspur and Raipur blocks may correspond to the Satpura strike in the rogion. 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.Tiku and Dipan Dehari Atomic Minerals Division. Departmenl of Atomic Energy. Begumpet, Hyderabad-500 016

An integrated geophysical approach was attempted for uranium exploration in Jamiri area. Arunachal Pradesh, using the techniques of magnetic, self-potential (SP) and resistivity profiling, coupled with solid state nuclear track detection (SSNTOJ, to (Icliiierite favourable structures controlling uranium mineralisation in phyllitic auartzites and quartzites of the Precambrian Dalinp fnrniiition.

Three promising zones of uranium mineralisation were recognised based on integrated results from these surveys. Magnetic survey identified llthologic contacts and faults in the area. A high-order SI' anomaly of -900 mV was observed near the contact of phyllitcs in the east and phyllitic quartziles in the west. A very low resistivity 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 characteristic of a fault that channelises radon and gives low resistivities. The SP anomaly may indicate sulphide mineralisntion and hence uranium mineralisation in this contact zone nidy be associated with sulphides.

The phyllitic quartzitcs wci»t of this contact are characterised by magnetic 'highs' 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 fX) •MMIIK «XHMiKtf

Figure 1 Geological imp of Che area wiCh uranium occurrences. 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 - phyllllLc cpjii rl/.lies rug I sic red u higher order oi susceptibility (10 to 20 x 10~6 cgs units) than quartzites (10 x 10 cgs units) including the samples j> 1 llu-d ll|> II'IIIII .iniiiii.i I nilr. r.'itl I II;II* I I vi- tint «• 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- wrslcru sltlc. In ^LMIL1 i"d 1 . L liu magnetic strike appcarb lo coincide with the regional 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 map showing anomalous magnetic zone, probable faults 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 magnetic profiJe (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 parC 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.

_ , . - J ? I Profile Profit* Qu I

A—j a-.,

Figure 6 : Downward cont lnuat Ion oC Sl» profiles (HP' and QQ1) - 59 - of SP. This fault clearly demarcates high gradient SP anomalies 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 SP along 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 phyllitic quartzites occuring within the zone L.are characterised by track density ano»aly closures of 60 to 200 2 tracks/.* . This zone falls in between tTie faults F1-F in the east and F2-F2 in the west. The track density anomalies nearby F1-F1 range between 60 and X20 tr«cks/mm2 «,rked as zone M which registers an SP anomaly of -900 -V with low resistivity of 1.0 oh* «. Similarly the track density anomalies near the fault

F2-F2 narked as zone N, are of the order 80 to 140 tracks/mn2 and colncide wttn nlgh sp gradlent UranluiJ mineralisation nay be associated with phyllitic quartzites in zone L, whereas the faults Fl-F1 and F2-F2 may be acting as conduits for the radon migration from depth. - 62 -

Although Arunachal Pradesh provides a heavy surface 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 fault F2"^2 is alao 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 and 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 - ground v.iluc, Lnd Lc;il <_• Lhul the |j|iy I I i L I u qua 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

1-0 Km =1 1 N OEX 1-T-ri BUXAS (OOLOMITIC UKSBNCfiRAPHITIC rT~n SLATES CALCAREOUS OUARIZ1TES)

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OMJNSS PHYLUTE OUARIZ1TE SEQUENCE

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|-"--'.| TENTATIVE CONTACT. - 67 -

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El LEGENO E3 E< ES EC E7 EB E9 RADIOACTIVE OUTCROP. - 68 -

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Contour interval 20 and 100 gammas E5 E6 C'7 E8 £9 ARBITRARY DATUM LEV EL = 47,000 GAMMAS - 69 -

Profile Mi urn

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80

i H» 1 Unil

I H>2U»it>

' H>) Unilt

Unilf • 0 - 70 -

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0 Traverse RA

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- 74 -

I I i:.-.Ti;:-fAi. T;i£vr;ci.Unii:-:-cy:c^ (i' "(.'HCLII-RCCK AM A X<.'I J:;:VJ. I.».I. TLU. Ill I'll! EX.-_C.'*ATIu\ ;. F iiASD£JTC:NE-rYi;H UIUMUi". iJ.^Ci Ai 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 Minerals Division, Department of Atomic Energy, 1 2 5 Ban:,alore-560 072, Hyderabnd-5OO 016, and ShillonE-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 Kopillis Formation Early Tertiary 1 Sheila Formation (alternating Formation coal-bearing sandstones and- limestones) Upper Cretaceous t Langpars (calcareous.sandstone) irA^,^^ Upper Mahedek sandstone Formation ^r Hahadek sandstone Jadukata conglomerate Jurassic » Sylhet Trap Precambrian t ijhillong Group metasediment3 Basement 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 MIS. MEGHALATA SHOWWG RABOACTIVE OCCURRENCES

^^^Si^^j ^k > |,-T

: ji I- I- I- ^^ f l» 1- ^ F' H^ - ==fc= 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,0fl Whole-rock Quartz-predominant mineral fraction

A. Bomiasiat area 1 0.015 * 215 and 255 185 and 235 2 0.039 'f> 215 and 270 190 and 240 3 0.011 * 210 and 255 185 and 240 4 0.058 * 210 and 255 185 and 235 5 0.14- % 210 and 270 190 and 220 6 48 ppra 250 235 B. Gomaghat area 7 0.033 * 205 and 255 185 and 225 8 0.061 5* 215 and 260 190 and 230 9 0.037 # 200 and 250 185 and 225 10 1.00 # 200 and 255 168 and 220 11 12 ppa 260 225 12 8 ppm 255 225 13 38 ppm 255 230 C. Pdengsbakap area 14 0.096 # 200 and 250 184 and 220 15 0.005 * 205 and 250 185 and 220 16 0.044 * 205 and 255 194 and 235 17 0.13 % 205 and 250 170 and 220 18 26 pp« 265 222 Uraniferous earn-I| 210°+ 10 180®!• 14°C and pies (* level) I 260°+ 10°c *n 230°+ 10wC Mean U-poor samples j| 260°+ 10°C 230°+ 10°C (ppm level) I

Table II. U,0Q, ThOp, and K contents of whole-rock and NTL peaks SI.No. ThO, K Whole-rock (°C) Quartz-riccmh portion 2 0.089* 0.006* 1.0* 215 and 270 190 and 240 6 48 ppra 37 ppm 1.0* 250 235 10 1.00 * 0.018* 200 and 255 168 and 220 11 12 ppm 19 ppm 1.0* 260 225 16 0.044* 0.015* 3.0* 205 and 255 194 and 235 17 0.13 * 0.034* 1.8* 205 and 250 170 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

Bhattacharyya, A.K., Rao, C.N. , and Kaul, I.K. (1976). Thermo- luminescent characteristic correlation and depositional environments of the Vempalle dolomites (Algonkian) in parts of A.P., India. Mod. Geol., v. 5, p. 237r254. Charlet, J.K. (1969). Utilisation des courbes de therrao- luminescence artificielle dans 1'etude des series sedimen- taires detritiques. Bulle. ooc. Geol. Fr., v. 11, p. 287. Charlet, J.K., Ducuis, Ch., and Quinif, Y. (1978). Mise en evidence par la thermoluminescence (i'L) des sables landeniens d1 anomalies radiornetrioues nouvellos dons la coupe du canal de Blaton. Ann. Soc. Gcol. Bel[> , v. 101, P. 537-349. Dhana Raju, R., Venkataraman B., and Anantharaman, K.B. (19S4). Natural thermoluminescence of whole-rock sanples as an aid in uranium exploration: A cese study from SinghbhUBiJshear zone,, ^ihar, India. Uranium, v. 1(3), p. 279-287. Dhana Raju, R., Selvam, A.?., Virnave. G.N., Sinha, R.M., Rajendra Singh, and Saraawat, A.C. (1989). Characterization of the Upper Cretaceous Lower Mahadek sandstone and its uranium mineralization in Dotniasiat-Goma^hat-tdenshakap area, Keghalaya. Explor. Hes. Atomic Minerals, v. 2, (in press). Gansruli, D.K. and Haul, I.K^ (196C). the a,^e of radioactive mincralisotion of placer deposits cf liorola, India. Econ. Geol.. v. 63, p. £)38-839. Hochman, K.B.M. and Ypma, >".J.M. (19S7). The accretionary migration of uranium in Tertiary sandstones - Thermolumi- nescence evidence from the fieverley deposit, South Australia. Uranium, v. 3, p. 245-259. Johnson, K.H. (1968). Determination cf magma temperatures from natural fcherrr.oluminescence. In: I-icDou^al (Editor), Thermoluminescer.ee of c;oolop;ical inatcrialc. Academic Press, London, p. :'•4-5-546. - 88 -

Kaul, I.K., Ganguli, U.K., and Hess, B.tf.H. (1972). Influr encing parameters in thermoluainescence. In: McDougal (Editor), Thermoluminescence of Quartz. Mod. Geol., v. 5, p. 201-207. Levy, P.W. (1977). New thermoluminescence technique for mineral exploration. Froc. IAEA Syrop. on 'Nuclear Techni- ques and Mineral Resources', Vienna, p. 523-538. Manconi, J.W. and KcDcugal, D.J.»(197O). Thermal activation energies of shocked and strained quartz. Amer. Mineral., v. 55, P. 398-402. McDougal, D.J. (1966). A study of the distribution of therno- luminescence around ore deposits. Econ. Geol., v. 61, p. 1090-1103. McDougal, D.J. (1968). Natural therraolurainescence of igneous rocks and associated ore-deposits. In: NcDougal (Editor), Thermoluminescence of geological materials. Academic Press, London, p. 527-544. Mukerji, S., Sengupta, S., and Kaul, I.K. (1981). Studies on some influencing parameters in the thermoluminescence of natural fluorites. Rod. Geol., v. 8, p. 1-11. Kambi, K.S.V., Bapat, V.N., and David, M. (1978), Geochrono- logy and prospecting of radioactive ores by their thermo- luminescence. Indian Jour. Earth Sci., v. 5» P» 154-160. Jiambi, K.S.V. and Mitra, S. (1978). Thermoluainescence investigations of old oarbonate sedimentary rocks. I'eues. Jahrb. Minor. Abh., v. 133, p. 210-226. Hishita, II., Hamilton, M., and Haug, It.M. (1974). Natural thernclumincscenco of soils, minerals and certain rocks. Lioil oci., v. 177* p. 211-219. larks Jr., J.tf. (1953). Use ..f Uhenr.oluuineacejice of liae- i-.tcno in subsurface o';riiti!iva.\-hy* Bulle. Auier. Ausoc. Pet. Geol., v. 371 p. 125-142. Sadaoivan, a., Uambi, K.3.V., and tiurali, A.V. (1981). Geo- cho.iicul ami Uijorino]uininocc«ncc :;tiKlic.'; ci' .•vijole:; froo ofi'-ahcre drill core, ..'e^t Coaet Ir.dia. Mod. Geol., v. 8, p. 1*-22. - 89 -

Sankaran, A.V.f Nambi, K.S.V., and Sunta, CM. (1982). Current status of thermoluminescence studies on minerals and rocks. EARC-1156, Bhabha Atomic Research Centre, Bombay. Sankaran, A.V., Nambi, K.S.V., ar.d Sunta, CM. (1983). Pro- gress of thermoluminescence research on Reological materials. Proc. Indian Natl. Sci. Acad., v. 49A, p. 18-112. Sankaran, A.V., Sunta, CM., Nambi, K.S.V., and Bapat, V.M. (1980). Therraoluninescence studies in geology. BARC-1060, ' Bhabha Atomic Research Centre, Bombay, 96 pp. Saunders, D.P. (1955). Thermolnminescence and surface correla- tion of limestone. Bulle. Araer. Asr.oc. Pet. Geol., v. 37» p. 114-124. Spirakis, C.S., Goldhaber, M.B., and Reynolds, R.L. (1977). Thermoluminescence of sand trains around a South Texas rollrtype deposit. U.S. Geol. Surv., Open-file Rep. 77-640, 14 pp. Vaz, J.E. and Sifontes, R.S. (1978). Radiometric survey using thermoluminescence dosimetry in Cerro Impacto (Venezuela) thorium deposit. Nod. Geol., v. 6, p. 147-152. Wintle, A.G. and Huntley, D.J. (1982). Thermoluminescence dating of sediments. Quart. Sci. 3ev., v. 1, p. 31« Zeller, E.J. (1954). Thermoluminescence of carbonate sedi- ments. In: Faul, H. (Editor), Nuclear Geology. Wiley, New York, II. Y., p. 180-188. Zeschke, G. (1963). Thermal RIOW tests as a Ruido to ore- deposit. Econ. Geol., v. 58, p. 800-803. - 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, 2 2 2+ viz., CO ", HC03", Cl~, SO4 ", and cations, vis., Ca , 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 =13 ppb). The U/Conductivity ratios for these shale and limestone unity r

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

The geological succession of the area (moditied King* 1872), mostly followed in our work is as follows:

1 t

1 i •

1 | KUNDAIR N/-iNDYAL OH/J1.E

1 GROUP KOILICUNTLA LIMESTONE

I KURNCOL PINNACLED QUARTZITS 1 PANEK

% SUPSR- GROUP PLATEAU QUARTZITE 1 GROUP 1 Middle JAMMALA OWK SHALE to • MADUGU N/iRJI LIME SI' OWE U.Prote- « GROUP rozoic • (1600- 600 m.y) , QUARTZITB & C0H5L0- GROUP

• I

1

i a UNCONFORMITY

1 i • , <\.I«jiNi* 3RISAILAM QU.iR'i'ZITES GROUP KOLUI^ULA SHALES IRL/JCONDA w'UAKT

' GUCUt N.vLLA- CUMBUM KALAI GUOUt BAIKEKKOHDA TjU, lUIV.IY

PULL. J1P£T/ T« 4UP. .TKI J VATHI H. .a .RI/PULVL'tlDI^» iUA GXCUP

P..PAOHNI lM^-rOHE & Gi.OUP JIL'-LE GULCHSaU QUAUTJIl'L'S

AKCHAEN JUL.-.;< CJN.i 3JE3 & ITi- /I'i'M ("T 2600 m.y) vji^ or Gu - 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,

2+ 2+ + + 2 Ca , Mg , Na , K , CC^~, HCO~ , Cl~, S04 " 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.

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

TABLE II, U. Conductivity and U/Conductlyity of well waters from the Cuddapah basin

CONDUCTIVITY U ppb (micro Mhos/Cm) U/Conductivity

UNIT n , Mean , Std. , Thre- , Mean , Std. Mean , Std. ' Thre- t Dev. , shold t Dev. , Dev.' shold

1.Basement • 20 73 83.6 240 1695 880 O.044 0.04O 0.124 Granite

2.Qulcheru 14 2.5 1.8 6 489 220 0.055 0.002 0.009 Ouartzite 3.Vempalle Limestone 113 8.5 7.6 23.7 910 435 0 007 0.005 0.018

4.Pulvendla Quartzite 71 4.0 3.7 11.4 650 260 0 .006 0.004 0.014

5.Tadpatri Shale 388 8.5 19.0 46.5 1500 1232 0 .006 0.008 0.022

6.Bairen- konda 9 3.4 1.8 7.0 1014 448 0 .003 0.001 0.005 Quartzite

7*Cumbum Shale 612 9.2 12 33 1201 467 0»007 0.006 0.012 S.Nandyal Shale 926 12.6 14.4 41.4 1829 1660 0.007 0.004 0.015

9.KoiI-Kuntla Limestone 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 Hydroqeochemical surveys

No.of anomalous Area No.of Mean Litho-Unit zones OCm2) samples U ppb deleneated analysed in well waters

Cuddaoahs: 1. TADP&TRI 1 30 28 44 SliALZS

2, CUMDUM 4 3O 54 18 SHALES

KUHNOOLS 1. NANDYAL 10 91 126 .54 SHALES

2. KOIL KUHTL,* LIMESTONE 1 18 32 35

i'Gi:..L 16 169 240 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 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. pT<»i«»aqi - i 1 H HI 7 "St a an M. ». • M 14 r. r. id HI i •» • nu •H Ml MM • . ai a it M • m • • - - — - • •• • • I - -

at •*« an a I- m - •HI a t M >. m -MB m. m • a> a** •141 M 11 r« •i 1 m m M • a - • - - • • - - — • -

t> -

••» HI - HI an «a» •n * • It * t ft "* n - aiwSS!* i l MM art an H 4 WIMI t n '-.

(BI*I M Hit a n a nwh

n •a HI M •n m 4 1 • « „ •t rt ft « .. «, t< a M • m wt M a • • — - -

Ml HI Ml It* •m ax •M u *>• *• r< n n t .- .. •i mi 411 .» n HI •n M in •> t M M 1 - M HI • « - - IP fffHHri n n ra n crnrn n c cr cr r Ka cr c: ^i ta HJ ca m ts ca E LulU - m nm n n ra i nm •« ai < r. MVMMW 1 IS t*" .a 1 u a I •»»

- £01 - - 104 -

77' 70" 79' _B0^ 91

MAP SHOWING CUDPAPAH 0ASIN A. P.

50 Scale 50 Km

11*

•Guntur

INDEX 16' DD KURMOOLS BO CUODAPAHS IV] NELLORE SCHIST BELT 5} GRANITIC GNEISSESES g Area covereco ered bv Ceochemlcal survey*

.15'

Nellora

M9

.13* 01' Flyure l H i ! • j>1 til I i s In hiii i ih i

I

O

I

I' - 106 -

COMPOSITE PLAN •r IMMMM *•• imaHOocinniTr CVHOUIS

... »»•- — n'. •<"••

HO •«—"«t Maou IMMMIlt IM MUM »••-.. ._ f >~| BIIWI H **' ~ |'.7_) IMMt MMMMlk ^6 MMtl 1 'H ;i£ i T

( ••) W—MUWHI MHMIMI ',] NT - 107 -

KURNOOLS. HISTOGRAM BASEO ON 'U' CONHNI IN WtU WAH R COMUIAIIVE FREQUCNCV CURVE SAMPLES BASEO 0N\)'CONILNI IN WflL 10 IN WEIL WATER SAM Pit II •0 *0 (0

in MnOf • **r* to 1=' i* to JO 10 10 f 70 •o

•M UO UO'IM MO ~IOO to •0 100 DO UO IM too

yEMPALLE DOLOMITE HISfOtlUM MX0 OM'U'MIUC* CUMULAflVC PERCEHFA&r VS'UTPk IN wen tMirn IN WEIL WATER}

u»»

Pijuro 4a - 108 -

TADPATRI SHALES

HI5TOCRAM BASER ON "If VALUFS CUMUlAtlVE PERCENTAGE VS'U'PPk IN WELL WATERS IN WELL WATERS inn

90

•0

70

60

So u

M

20 10LIL^ 7* M W M M «9i»AMItE

CUMBUM SHALES

HISTOGRAM BASED ON V VALUES CUMULATIVE »CMCENUK VSVPfk IN WCU IKATCOS IN WCLt WATERS

•e 7» _, '•CWfKf M •r.ctM*>

if

• MM M Ut IU IM M 1M

Figure 4b - 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 mineralisation becomes difficult in areas uhere soil cover is considerable, because of poor gamma ray response. In sycb areas, measurement of integrated concentration of soil gas ^zRn along uith cumulative gamma dose helps to detect the concealed uranium mineralisation. This combination method uas tried in Singhbhum Thrust Belt in eastern India uhere hidden uranium mineralisation uas suspected. Exposure cups equipped uith cellulose nitrate films used as detectors for measuring soil-gas radon concentra- tion and CaS04 (Dy) thermoluminescent dosimeter for measurement of 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 226R3.

* 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.qusrtz congiomarata Iron ore stage - phylllCes, Iron Ure Stsgu - phylllta,tuff, quartzites, arkorfe, tuff and baalc conglomeratet ignaoua rocka quartzita and baeic Ignaoua Chalbaea ataga - ailcs achlat rocka. hornblanda achlat quart achlat and tuffa. - 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 of radium, a member of the 23flU decay series, which is widely dieti butad in tha earth1a crust. Radon ia an inert radioactive gat, - 112 -

which decays with a mean Ufa 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 ax 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 Mia waa normalised to 30 days exposura and than converted to radon axpoaura uaing tha calibration aquations

0#98 CE - 20.08 x T 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 and t 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

Table - I

Radon concentration and Integrated gamma values in soil-gas of Rajdoha - Dunqrldih

Gmmtrm dos« 222 222 Sample Sample Gamma dose ftn cone. •illi re»x30d) No. ( Rn concn. No. (nilli rexx30d) (pCi/1) (pCi/1) 1. 90.40 986.70 23. 1007.00 1202.80 2. 90.70 756.60 24. 878.00 130.60 3. 133.00 1257.00 2S. 771.00 100.40 4. 191.00 364.70 26. 636.00 475.80 5. 150.70 169.70 27. 933.00 114.80 6. 134.30 501.00 28. 877.00 58.20 7. 102.10 460.30 29. 860.00 146.30 8. 105.70 596.30 30. 833.00 36.10 9. 112.80 248.80 31. 624.00 68.20 10. 116.40 401.10 32. 731.00 530.00 11. 107.10 292.50 33. 578.00 350.70 12. 192.10 400.00 34. 695.00 362.90 13. 181.40 486.00 35. 703.00 455.40 13b. 857.00 1014.10 36. 810.00 102.30 14. 202.10 SS5.60 37. 692.00 699.00 15. 130.00 806.80 38. 692.00 201.00 16. 1057.00 404.60 39. 675.00 236.00 17. 908.00 1167.00 40. 630.00 654.80 18. 1034.00 1117.60 41. 559.00 264.00 / 19. 903.00 1373.70 418. 480.00 1114.60' 20. 981.00 1351.40 21. - - 22. 953.00 1891.20 - 119 -

Statistical evaluation of aoll-qaa

222 Rn concentration data

Table - II

Location Statlaticai information fig C Porcentile (pCi/l) 95 60 20

ROHINBCRA 87.4 2.8 17.6 67.4 208.1

3AOUGUUA 311.1 2.4 79.3 249.5 651.5

OUNCRIOIH 263.7 2.7 56.6 205.5 605.7

RAJDOHA 602.4 2.0 203.90 505.5 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 UlT. - 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 £

o

6

a - in 2-

% •

* • J a ru 3

I X • t

* UJ

t • I

i » i'

t - 124 - ff.*T «*.» «.» *• »» N TO «0 tO «0 M tO IS 1 1 I OS Ol !| MS

CUMULATIVE PERCENT MORE THAN STATED LOG PROBABUTY PLOT OF SOIL-GAS 222R«. CONC. OBTAINED FROM JADUGLOA.DUNGROH - RAJPQHA. • OUNCRCM • RAJOOHA - 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 India Ltd.

Bhatin is a small uranium deposit being worked by UCIL in Singhbhua district of Bihar* Large dispersion of R.O.N. grades have been causing a concern to a great deal from quite some time. Conventional technique adopted for rwrvs evaluation lacked accurate prediction of grade fluctuations. Geostatistical technique is used for reserves evaluation. Error involved in estimation is calculated. Attempt has been made to study these wide variations in predicted and achieved grades. Estimated grades of the deposit by both the techniques are compared, best estimator for grade of unknown mining block is 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 grade is determined by logging of the boreholes and radlometric measurements and chemical assay of core samples of each bore- hole and the volume of the ore body,grade,accumulation (grade times the thickness) and other essential parameters are roughly delineated. Using these values, so far, it was customay to use classical statistical techniques(for example polygonal or inverse distance method and Slchel's *t* estimator) to obtain ore 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 repsatlaf the 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 experimental variocraaa 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 shapee and its equatioqn 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 coefficients of the sample grade gi

^ ~ (condition for unbiased estimation) -

^'O'?)/ *i ft* v) are *ne average values obtained froa 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 not possible, the equations are solved to get the w^ghting coefficients by suitable computer programmes and then the block grade and associated estimation variance are calculated. In the case of uranium where radiometric logs are used, cokriging can be made between chemical and radiometric grades.

Case Study ; fihatin uranium deposit is considered to be a structurally controlled hydrothermal deposit. The principal structural controls are the shear planes which strike approximately NV - SE and dip to the NE. There are two lodes (1) the main or the hang wall lode which extends from east to west for a distance of 400 metres and (2) the foot wall lode which is much shorter(approximately 150 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 statistical 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 is of the transitive type and the continuity is maximum along strike and least across it. Strike variogram of bore bolt data depicts insufficient grid size used for evaluation of the nawvea of the* deposit. Waste zone is found to be increasing at the time of mine development. - 133 -

(5) Kriged estimate is found to be the best estimate for underground mining blocks. It has helped to control large R.O.M.grade fluctuations.

Acknowledgements ;

Authors are indebted to Shrl S.Shastry, Additional SuperiTtandent(Otology),UCTL, for his valuable suggestions and encouragement in carrying out this worK.

Thanks are also due to Shri J.L.Btiasin, Chairman and Managing Director, UCIL, for his support and interest for this type of work and bis kind permi- ssion to present this paper at this symposium. - 134 -

References

BLAISE, R.A. and CARLIER, P.A.,(1968) Application of Geo-statistics in Ore Evaluation, Canadian Inst.of flin. & 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 foundations of geostatistics and the practical methods of constructing a ssnivariogram, E & MJ, Vol.180,Mo,7,pp.90.

CLARK, I.,(1979) How to fit a sioplistic model to an experimental 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 TO STATISTICAL ANALYSIS, Hograv Hill BOOK CO., N«W Yor*.

JOUftifcL, A.G.,(i979)G«ost8tistica& Siaaulatlon methods for Exploration and nine Planning E A MJ, Vol.180,No.12,pp.86. PARKER, H., (1979)The volume variance relationship : A usual Tool for nine Planning, E A rtJ, Vol.180, No. 10, pp 108. RENDU, J.n.,(1980) A ease study ot kriging for Ore valuation and nine Planning, E&HJ 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 Gsostatistics to 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 .

X GO

I : SO

r,<

• A -i r s < 3! I CO * ( 1

••-, .:.-•?

-"• . \. —.•;• *~

»'.• .«. • r_> '.•rtt.- :••• 137 -

...i"

LOG-ASSAY-FREQUENCY HSTOGRAM- OF B-H-DATA

100 NUMBER OF SAMPLES *» 417 .. ,

SO 1 • .

A i . 1 \ . I .... .

40 ' ' ' ''i' '

V • •• ;: : : .' • . I .'. i '• -:'.';... : '' K>

0 W . M- 12 M M 15 Ifi l> 18 13 20 . ?l r 2-3 24

LOG Af.SAV ->- - 138 -

——-f

f QT •_= - v -r,.-i---g&ar-=-.:..,H_-, j;^, 4. i..

TOTAL No OF JS^MPLES '417 -

• . • • •••••: y -•;• ' - -K-O-

•iSO - • : • / -— - —

• •• 9 ISO

•tic I

•- •- —HO~i (

•:*?! i I3C- t 9 -• *? •no uo

< < -C70

•MO

osa

.-- -CiO Fiq.3 -020

•01? - 139 -

r!

_ j.c.

t.0 • *6

-O--H ::_/

'•3

i-c.

.— J_^_:_ CUMULATIVE/-* i.

PLOTJOF ^UHULAra>"E PPEQUEKCY PERCENTAGE Vs LOG ASSAY

"TOTAL No-OP" SAMPLES = 417 •"•""• : - uo -

\

c°

C O o o

5 '-^

CM'

• ORE HOL.E LOCATIOM • HAT IN 'MINI SCalfti- V. 4OOO Fi

- 142 -

• _1. •• I

SEMI \»RK)GRAM OP ACCUMULATION VALUES / THE STRIKE ..

OF ORE BODY C+d5m LEVEL)

. No Of BOREHOLES =3

No OF SAMPLES* 3

XCJUO

2MXV0

26OHG'

240X19*

it 143* |C5

JO

40II10*

20X10

KC IcC W if. DISTANCE IN METRES-> F/j.7 - U3 -

ACCUMULATION FREQUENCY ' HISTOGRAM OF CHANNEL DATA NUMBER OF SAMPLES = 84

tc

i i 12

2

2

O 1 1 0 2 4C 12 Id 16 16 ' 20 22 24

ACCUMULATION - 144 -

LOG - ACCUMULATION FREQUENCY HISTOGRAM OR -.CHjANNEL

20 NUMBB'liQF, SAMPLES_-

M i i ; •' .1 . .-.:.•. v- M :.1:;ll

— IUl it--

I 1 JZ1__. ..:«6 »S0 3-80 400 405 4-15 . 485. 4-30 i

.• * LOG -ACCUMULATION - 145 -

SI"?1 \ I'M j; '• >

• ••

w.-r"

. o

1 t ; ,:.-..• j

40 u>o no ieo wo 200

I!! ME.TRL-S • •-

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I a:

I a 9 Q: SESSION II B

ANALYTICAL TECHNIQUES IN URANIUM TECHNOLOGY - I

Chairman : Dr. tf.C. JAIN, BARC Reporteur: Shri V.N. KRISHNAN, BARC. - 147 -

URANIUM ANALYSIS USING AN ON-LINE BACKGROUND CORRECTION PROGRAM WITH CARRIER DISTILLATION TECHNIQUE BY A COMPUTER CONTROLLED EMISSION SPECTROMETER

R.K.Dhurawad, A.B.Patwardhan. V.T.Kulkarni, K.Radhakrishnan Fuel Reprocessing Division, B.A.R.C., Trombay Bombay - 400 085.

SUMMARY The paper describes an on-line background correction and uraniua monitoring (due to occasional matrix excitation)in a computer controlled Direct Reading Spectrometer during the estimation of a large number of impurities in uranium product used in nuclear facilities. The influence/interference of the background on the analytical lines becomes important when low detection limits are to be achieved. Commercially available softwares for automatic background correction (ABC) are suitable for Inductively Coupled Plasma or Spark sources where one can have a continuous, flow of the sample introduction (without much restriction on time of exposure). However, in the case of carrier distillation technique automatic background correction cannot be applied due to limitations of exposure per charge. In the method discussed here, a background channel located at an appropriate position in the spectral range is monitored simultaneously along with other analyte channels. The background at analyte channel is computed from the intensity of the background channel and is automatically subtracted from the intensity of the analyte signal. In addition, a uranium channel which is used for Monitoring a weak line of uraniua (286.567 nm ) is incorporated to measure the amount of uranium getting into the arc. When intensity of uranium line exceeds the predetermined value, the data will be rejected by the operator. This method is in routine use over a few years for the estimation of 2t impurities in uranium. INTRODUCTION

For quality control of uranium required in nuclear facilities, carrier distillation technique was developed by Scribner 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 optimised conditions. The limits of detection for a majority of elements are in the range of a few parts per million. For B and Cd it is necessary to ensure that the detection limits are 0.1 ppm or better. The influence/interference of background on analytical lines becomes especially important at these levels.

Until recently the impurity analysis was carried out in our laboratory using photographic imaging and densitometry. It was a time consuming process, with the advent of new technology i.e photo-multipliers, microprocessors and computers, the photographic detection has been replaced by Direct Reading Spectrometers. Pepper & Blank (2) have reported use of Direct Reader for carrier distillation using exposure control unit for different elements. Background correction in Direct Reading Spectrometers for spark and ICP sources have been reported in literature (3,4) using two methods namely refractor plate technique and moving slit technique. Both these techniques essentially require a steady and continuous source over a total period of exposure. This requirement can be met in the case of ICP as well as spark sources. However, in the case of carrier distillation technique the sample is not continuously injected at constant rate but instead a limited quantity of sample blended with carrier loaded on the electrode is consumed during a short period of a fraction of a minute. Moreover different elements have different volatilization characteristics depending upon the nature of the sample, atmosphere and electrode material etc. Thus the limitations of the sample amount and exposure time in the case of carrier distillation method are some of the major factors to be taken into consideration. It is not possible to apply the same method of background correction as employed in ICP or spark.

In some laboratories background correction in carrier distillation technique is carried out by subtracting known signal which is arrived at by measuring the signal of pure matrix at all the channels and then computing the background corrected intensity (BCD for each channel (element). This is somewhat an arbitrary subtraction.

BCI « I - I t element blank In the present method, the approach is unique and is totally different which can be termed as on-line background correction with Direct Readiang Spectrometer. In this approach, a background channel is located at an appropriate position in the spectral range. Intensity is read at this channel alongwith other e alytical channels. This intensity is used to compute the background of different analytical lines by multiplying intensity of background channel with pre-determined factors. BCI • I - I • R Elem.channel Dkg. channel Where *R* is pre-determined factor for different analytical lines which is calculated as explained below. In a given line, if IBL and IBR are the background intensities to the left and to the right of the line respectively then the average background is represented by Average Background - (IBL + IBR) / 2 - 149 -

Por each analytical line K is calculated using the formula ( IBL + IBR ) / 2 K • ————————————————————~ i Background channel IBL and IBR are found by reading intensity of pure matrix by moving the entrance slit of the spectrometer. RESULTS & DISCUSSION

Table I gives the observed and computed background intensities of a few selected channels .Thus incorporating all the constants evaluated for different elements in the equation, uranium samples were analysed. Table II gives the analysis results of two synthetic standard samples. The results are in good agreement.

The degree to which uranium interference is avoided is also an important factor in carrier distillation technique. To keep strict control on exposure, we have adapted a system wherein a channel for a weak uranium line (286.567 nm ) is monitored alongwith analyta lines and its intensity is measured. The cut- off intensity is pre-determined which is 1000 counts. Any exposure exceeding this intensity is rejected by the operator. The results of the exposures exceeding this intensity limit have been 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 of uranium samples. ACKNOWLEDGEMENT! The authors wish to thank Shri A.M. Prasad, Director, Reprocessing Group and Shri H.K. Rao, Head, Fuel Reprocessing Division for their keen interest and encouragement during the course 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,

TABLE I Observed and computed background intensities observed computed observed computed

Element B Element Cd 455 419 293 267 422 414 257 248 434 389 273 255 452 478 311 265 447 437 3tl 257 517 437 347 250 553 497 407 31* 662 513 399 308 696 527 387 298 559 531 365 29t

Element : Co El •suit : Zn

1199 1443 386 377 1381 134* 359 350 1578 1377 369 359 1529 1434 369 374 1393 1387 404 362 1159 11M 472 352 nee 1075 385 362 1280 Ilt6 510 347 1302 1151 417 405 1122 1113 417 412 - 151 -

TABLE II (All values in ppm on uranium basis)

Sample 1 Sample 2*

Element Amount Amount Amount Amount added detected added detected

Al It 6.8 — B 5 4.9 0.1 0.13 CO 2 16.7 1 2 Ca 10 14.7 - - Cd 20 25 0.1 0.16 Cv 10 9.1 10 8.10 Cu 10 9.6 2 2.0 Pa 50 50.8 - - MO 25 2% - - Mn 1* 9.2 1 1.3 Ni 5 4.2 10 9.9 Pb It 9.5 2 1.5 Zn 20 22

* Sample 2 is a R«f. standard wherein some of tha impuritias like Al,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 3t psi Sampling depth s 23 mm Nebuliser argon flow 0.4 lpsi Measurement time : 0.25 sec Coolant argon flow 12 1pm Measurement/peak : 3 Auxiliary argon flow 2.4 lpm Repeat integration : 8 Plasma power 1.2 Kw Reflected power 5 Watts - 155 -

Table 2. Matrix effect of uranium. Cone: Percentage suppression of the analytical signal of uranium Li Cu In Ce Ho Bi 1000 ppm 86 86 75 70 65 66 500 ppm 67 59 45 46 38 25 200 ppm 40 29 6 12 -24 100 ppm * * * -24 -30 -50 * - No suppression. Table 3. Direct nethod Element Con.in RSD y. IAEA Del:. limit FBR Speci IAEA Certified (ppm) -fications sample range (ppm) limits (ppm) Ag <0. 9 _ _ 0.9 20 Cd <1.1 — — 1.1 1 Cr 5.9 12 3. 1 - 5.0 2.4 300 Co 4.3 6 4.0 - 4.3 0.5 200 Cu 8.6 13 4. 3 - 6.7 3.6 100 In <0. 5 — 0.5 Mg <4. 9 — — 4.9 150 Mn 14.3 7 14.0 - 18.0 0.6 200 Mo 11.4 3 9.5 - 16.8 1.0 200 Ni 13.8 5 8.4 -14.1 3.1 500 Sr <0. 4 0.4 150 Ti <1.9 — — 1.9 100 Elemiint Det •limits El em»nt Det.1imitii Ce 0 .3 Gd 1.8 Er 0.6 Pr 0.1 Dy 0.7 Sm 0.2 La 0.5 Ho 0.4 Nd 0.6 Yb 1.5 Table 4. Solvent extraction method ElNMWit Con:i n RSD X IAEA IAEA Det. IAEA overa1) Certified 1 imite •ample median range

References:

1. D.J.Douglas and R.S.Houk.. Inductively coupled plasma^ mass spectrometry. Prog.Analyt.atom.spectros. Vol-8. pp 1-18. 1985.

2. G.M.Hieftjc, and G.H.Vickers. Developements in plasma source/mass spectrometry. Analytica Chimica Acta. 216. pp 1-2-4. 1989.

3. A.G.Pa«e etal.. Estimation of metallic impurities in uranium by carrier distillation method. BARC-862. 1976.

4. A.G.I.Dalvl et al. Chemical separation and spectrofraphic estimation of rare earths in (UPu)02: Talanta. Vol.24, pp 43-45. 1977. z

5. B.D.Joshi et al.. Anion exchange separation and sttoctrofraph1c determination of rmrm earths in lutonioum with LiP/AgCl carrier. Anal.Chim. Acta, f7. 379-86. 1971. t 6. T.R.Banfia et al. Spectrochemical determination of trace metals in uranium. B.A.R.C - 950. 197S - 157 -

DEVELOPMENT OF FLOW INJECTION ANALYSIS TECHNIQUE FOR URANIUM ESTIMATION A.H. PARANJAPE; S.S. PANDIT; S.S. SHINDE; A. RAMANUJAM; R.K.DHUMWAD Fuel Reprocessing Division BARC,Bombay 400 085 Flow injection analysis is increasingly used as a process control analytical technique in" many industries. This paper describes the development of such a system for the analysis of uranium (VI) and (IV) and its gross gamma activity. It is amenable for on-line or off-line monitoring of uranium and it3 activity in process streams. The sample injection port is suitable for automated injection of radioactive samples. The peformance of the system has been tested for the colorimetric response of U(VI) samples at 410nm in the range 40 to 350mg/ml and U(IV) samples at 650nm in the range 15-120mg/ml in nitric acid medium using Metrohm 662 Photometer and a recorder as detector assembly. This technique with certain modifications is used for the analysis of U(VI) in the range 0.5-4mg/aliq. by alcoholic thiocynate procedure. In all these cases the precision obtained was found to be better than +/-1.5X. With Nal well- type detector in the flow line, the gross gamma counting of the solution under flow is found to be within a precision of +/- 5%

I. INTRODUCTION Flow injection analysis(FIA) is a simple and elegant technique which finds increasing applications as a process control analytical technique in many industries. Ruzicka and Hansen were the first to use the term Flow injection analysis (1)). In general it involves injection of a sample aliquot into a steady flowing stream of reagent and passing this reagent- sample mixture through a suitable detector. FIA is a flexible and convenient technique which can be adapted for continuous process stream monitoring with good precision. In a laboratory environment, such a system can handle many samples vmry quickly and because of the flexibility, the same system can be modified to carry out various analyses. This paper describes the development of such a system/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 of uranium extraction and purification plants and in fuel reprocesing plants based on Purex process. The sample injection ports are suitable for automated injection of radio aotive samples II. REAGENTS AND CHEMICALS Uranyl nitrate solution in 0.1M nitrio acid and electrolytically generated uranous nitrate in 1.0M nitric acid - 158 -

and 0.1M hydraaine were used and their concetrations were estimated using standard analytical procedures. Alcoholic thiocynate reagent containing stannous chloride and ethylacetate was prepared as described elsewhere(2). III. DEVELOPMENT OF FLOW INJECTION ANALYSIS SYSTEM The technique for the estimation of U(VI) in the range 40 to 350 g/1 involved injection of the sample aliquot at a steady rate into a steady and continuously flowing stream of dilute nitric acid and measurement of its colorimetric response at 410nm with a suitable photometer. For monitoring the gross gamma activity, the same diluted solution was passed through a gamma counter. The modular configurations used are given in Fig.l. With minor modifications in the sample delivery unit and injection port assembly, this technique could be used for on-line or off- line analysis of uranium. The method used for the analysis of uranium in the range 0.5mg to 4mg was based on colorimetry at 420nm, with alcoholic thiocynate as the chromogenic reagent. To carry 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 at the injection port or circulated in a closed loop till a steady state response is achieved at the detector. Though any of the above two configurations could be used for the colour measurement of U(IV) at 650nm after dilution with nitric acid, the tirst configuration was tested.

The FIA system can be divided into three modules: a) delivery units for maintaining steady flow of the reagent ( and sample, if required) b) injection module for introducing sample and c) detector module. The various units employed in the above mentioned configurations are detailed below.

(a) Reagent Delivery Units For direct colorimetric estimation of uranium, a microprocessor controlled Multi Dosimat Unit (Metrohm, Swiss, model No. 665) was used to deliver 0.1M nitric acid at a constant rate that can be adjusted to any desired value. This unit was used with 10ml volume burette (No.6.3007.210). The accuracy of the volume delivered is +/-0.2X. For delivering the alcoholic thiocynate roagent, a piston pipette (Repipet.USA ) with a capacity of 10ml/stroke was used for delivering 10 ml reagent at a time with a volume accuracy of +/-1.5*.

(b) Sample Delivery Units The sample aliquots were delivered by another Multi Dosimat with one ml capacity burette (no.6.3006.113).With this unit, any aliquot within one ml could be delivered at a constant rate that can be adjusted. The accuracy of the volume delivery is +/-0.3X. For on-line monitoring purpose with a dedicated FIA system incorporating the Multi Dosimat Unit, the sample solution was - 150 -

sucked straight into the one ml burette and delivered Into the injection port assembly via a permanent microbore tube connection and this operation could be repeated any number of tiroes.

For handling radioactive samples during off-line analysis of uranium, the delivery mode of the Dosimat was modified to avoid cross contamination. In this case, the burette was connected directly ( without, going through the three way valve) to a disposable polypropylene pipette tip of one ml capacity through a flexible narrow bore teflon tubing of appropriate length ( 1 ml capacity). The burette and the teflon tubing were filled with water such that during sample sucking and delivery, the water moves back and forth within the teflon tubing thus reducing the volume of the air pocket above the sample in the tip. The sample (lml) sucked into the tip was delivered into the desired injection chamber either as a continuous stream or as distinct aliquots of smaller sizes at the chosen flow rate. In the later case, usually the first aliquot was rejected. After pipetting, the tip was disposed off. The delivery volumes have an accuracy better than +/- 1.0%. The derails and reliability of a similar unit are described elsewhere (4).

(c) Injection Port Assembly

Injection port assemblies form a crucial part of the FIA system and in the present work, they had to be compatible with safe radioactive practices. Two different types of injection chambers as shown in Fig.l, were chosen for testing. Of these, the first one was a fully closed system meant mainly for on-line analytical applications and the second one was an op«m-cup type suited for automated off-line analysis in laboratories. (i) Fully Closed Injection Chamber The fully closed injection chamber assembly was made up of a pyrogen device used for glucose drip adjustment in hospitals, with additional penetrations for sample entry via microbore tubing and a vent line for adjusting the pressure build up and for controlling the liquid level inside the injection chamber. Shortly after starting the reagent flow, the sample was injected into the chamber at a steady rate where it got diluted and transported via narrow bore tubing to the photometer. The reagent and sample delivery points in the injection chamber were located in such a manner that during their delivery a good miking of both takes place in the chamber Itself. Being a closed system, the sample-reagent mixture passed through the narrow bore tubing under slight pressure exerted by the Multidosimat units. Sufficient tubing length (90cm long/1 mm bore) Was provided to get adequate mixing before the mixture reached the photometer.

For on-line monitoring, the sample delivery tube of the Dosimat was directly connected to the injection dhamber. Whenever this injection port assembly was used for off-line analysis, the - 160 -

sample was injected through the septum in the port using the polypropylene tip. (ii) Open-Cup Injection Chamber Inorder to avoid the necessity of injecting the sample through the septum during off-line analysis, an open cup was used as injection chamber. This had the advantage of easy introduction of sample with minimum cross contamination and is suited for automated sample addition. For diroct estimation of uranium, the sample was added into the cup using polypropylene tip at a steady rate and mixed with steady flowing stream of 0.1M nitric acid with the help of a magnetic stirring bar and the mixed solution was passed through the photometer. A minimum but constant liquid level was maintained in the injection cup by adjusting the T at the discharge point of the tubing to the same level. The main drawback with this system was that the tubing diameter should be large enough to allow the free flow of the solution at the same rate at which it was being delivered in to the cup. Further, complete mixing of sample and acid should take place in the cup itself.

For uranium assay by alcoholic thiocynate method using the stopped flo'w technique, the sample was added to a fixed volume of the reagent in the open cup and after mixing, the entire solution was drained through the photometer for absorbance measurement. As an extension of this method, following the closed loop procedure, the solution can be recirculated between the cup and the photometer using a peristaltic pump till constancy in the optical density is achieved and after which, the solution can be drained off. These two techniques will be useful for all spectrophotometric procedures that require some tine for colour development. In the present work, only the stopped flow technique was tested.

(d) Detector Modules For absorbance measurements a photometer (Metrohm Model 662) was used. It was equipped with a flexible, sheathed optical fibre light guide as detector. Normally this device is to be dipped in solution for absorbance measurement and is used mainly for end point detection in titrimetric analysis. However in the present instance, a pyrex glass tube of 2 mm inner dla was inserted in the light path of the detector tip and the solution to be monitored was passed through this tube. Thus the photometrio detector never came in direct contact with corossive and radioactive solution. Absorbance Measurements could be made at any desired wavelength in the visible range. After initialising the unit to sero with blank reagent, the absorbartee of the sample-reagent mixture flowing through the tube was - 161 -

monitored as digital output and the same could be plotted using a chart 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 was passed through a coiled tubing inside the well. The counting was done for a fixed time starting from sample introduction. The gross gamma count rates obtained were subtracted for the background obtained by passing the blank solution for the same time interval. IV. RESULTS AND DISCUSSION

Except in the case of uranium (VI) estimation by alcoholic thiocynate method ( where sample and reagent were mixed under static condition), in all other cases, both reagent as well as rample were introduced at selected flow rates. The colorimetric response of the FIA system as a function of sample and reagent flow rate was studied in detail to arrive tvi optimum conditions as it is an important factor in FIA analysis. For this study, estimation o* ••-•»•«•• nm bv direct, colori™*"' w«a used as a re/erence mecnoa.

The variations xu v-~.~ ...:ic response or the system «d a function of reagent flowrate at constant sample flowrate and vice versa are shown in Fig: 2a, b & c. These data were useful in optimizing the sample and reagent flow rates while standardising colorimetric procedures. Fig.2 a shows the variation in absorbance as a function of aliquot size for a given sample and reagent flow rate. It is seen that for aliquot sizes above 0.5ml, a constant response is observed at the detector, irrespective of the aliqout size. Thus above this minimum aliquot size, the FIA response becomes independent of the aliquot size aa long as the flow rates are kept constant. This is of groat advantage while planning online analytical procedures that require dilution or reagent addition before colour measurement*.

Table I-A shows the results obtained for direct estimation of uranium in the range 40-350 g/1 using closed chamber and open cup as sample injection assemblies. As the aliquot size was 0.5ml, a steady response for a reasonable length of time could be observed at the detector. It is seen that the precision observed for open cup injection is better than that obtained using closed injection chamber. This may be due to the introduction of sample through the septum in the later case which is not as reproducible as sample delivery in open cup. As the solution in this case is flowing under slight pressure it was possible to pass this sample reagent mixture through gamma counter (well-type). The cross gamma count rate obtained could be correlated to its concentration of uranium with a precision of +/-5X and these results are included in Table I-A. Similar colorimetric results obtained 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 two concentrations of uranium when the samples were injected into the closed injection chamber via permanently connected tubing(Figlc). The precision obtained in this case indicates that this mode can be used for on -line monitoring tasks where gradual variation or steady state concentration of uranium is to be monitored. Sample introduction using a six way valve is currently under investigation for the same purpose. Results obtained for the estimation of uranium in the range 0.5 to 4 mg per aliquot by alcoholic thiocynate method using stopped flow technique gave a precision varying from +/~ 1.5% to +/- 0.25% corresponding to the lower and upper limits of the uranium range tested. This precision is satisfactory for process control analytical applications. As aliquot sizes were small in these cases (0.05 to 0.2ml), these could be varied depending on the concentration of uranium without causing any major error.Thus this technique is useful for a variety of analytical methods that require vigorous mixing before colour measurement. As mentioned earlier, the same results can be obtained using closed loop flow technique. In this later case, a six way valve may be more useful for fixed volume addition, if cross contamination is not a major factor. V. ACKNOWLEDGEMENTS The authors wish to express their sincere thanks to Shri A.N. Prasad, Director, Reprocessing Group and Shri M.K. Rao, Head. Fuel Reprocessing Division for their keen Interest in the work. 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 and Radiation 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/min Sample aliquot 0.5ml fixed A. DIRECT COLORIMETRY OF U(VI) AT 410nm

Sample Closed-cup Open-cup Closed-cup

U O.D % RSD O.D XR.S.D Gamma counts «/l 100 sec.

350.0 0.619 0.62 0.589 0.51 3726 175.0 0.310 0.87 0.306 0.18 2046 116.7 0.216 1.58 0.210 0.54 1410 87.5 0.164 0.53 0.150 0.63 1037 43.8 0.083 1.07 0.083 0.62 558

350.0 * 0.625 0.45 - - - • 175.0 * 0.338 0.83 - - * Saaple line directly connected to closed cup B. DIRECT COLORIHITRY OF U(IV) AT 650n«

Closed cup Open cup U(IV) O.D % R.S.D U(IV) O.D XR.S.D f/1

118.0 0.766 0.60 104.0 0.705 0.30 59.0 0.432 1.10 52.0 0.356 0.33 39.5 0.300 1.50 34.7 0.262 0.29

29.5 0J229 0.66 26.0 0.214 0.57 14.8 0.125 0.68 13.0 0.111 0.10 - 164 -

SCHEMATIC DIAGRAM FOR FIA

REAGENT FLOW SAMPLE - INJECTION DETECTOR

TUBE MAGNETIC 0AR PINCH CH6ITAL DISPLAY OR RECORDER (a) OPEN CUP SYSTEM

REAOCNT rvom

DNMTAL OMPLAY aM&vunw—* ocncTow (t) CLOSED CUP SYSTEM WITH SEPTUM

MfAtfNT POJHWAT

FROM MMPLC OOWMAT—*•

OMtTM. OUTLAY OR HCCMOCM

OCTCCTOI (c) CLOSED CUP SYSTEM (WITH OUT SEPTUM)ON LINE

FIG -a - 165 -

EFFECT OF SAMPLE SIZE AT EFFECT OF VARIAtLE REAGENT EFFECT OF VARIABLE SAMPLE CONSTANT REAGENT FLOW FLOW AT CONSTANT SAMPLE SIZE RATE AT CONSTANT REAGENT CHAUT SHED :• ISM/M* AND CONSTANT SAMPLE FLOW FLOW MJOUOT (SOO.U WAVE LEMTM:- 410 *•> ALMUOT (BOOA) CONSTAKT ' WMn.t>l7S«/l U{V1) SAMPLE SKCD- Iml/m* •MWltXTM/l U(VI| KCMENT PLOW HATE-••('mill ItNKI FLOW MTCV-I«l/Ml« •AMP1E PLOW MTC VAMCO. ' *C«KMT S#H0 • «•!/•(• A — I mi/ ml* MMCNT PLOW RATE VMIMCO. SAMPLE MOt'l UIVII • m A> IM/tlhl • * 4«l/ •*«« C « C * tmi/mit, D •

c as* 0.M* A 0.478

0.300

•ooJk aoox 400JL sooA tml TIME ELAMCO

FIG. 2 - 166 -

STANDARDIZATION OP A D.C. ARC CARRIER - DISTILLATION PROCEDURE ON A DIRECT 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 powder J 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 J mixture. The AgCl contained 2Z Ca.O-. A 'BLANK' was also prepared, consisting of U 0 used in the J 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 of the analyte in 'BLANK*

Total grbss intensity of the Internal - Total gross intensity of standard 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 added Standard 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 Jobin-Yvon Model No. JY-48 Poly- chromator with lm concave grating in Paschen-Runge mount. Grating type, no. of grooves holographic, 2550 gr/mm Wavelength region, order 130-415 nm, I order Dispersion 0.39 nm/mm in I order Slit width 30 microns Analytical gap 4 mm Excitation source Stabilized d.c. arc operated at 8 amp. Exposure time 35 sec. TOTAL Electrode assembly Lower electrode (anode) Vdia. U.C.C. 1990 graphite elect- rode containing 120 mg of sample/ standard. Upper electrode (cathode) 3/i't dia. U.C.C. pointed graphite electrode. Data acquisition and through APPLE-IZe Computer. processing.

TABLE III Analytical Data

Mean RSD

Element Concentration range

B 249.7!) 0.1 - 1.0 5.9 10 Cd 228.80 0.1 - 1.0 8.4 9 Ni 231.60 10 - 100 10.8 10 Mn 257.60 1 - 10 11.5 13 Cr 267.70 10 - 100 9.2 11 Co 243.20 1 - 10 6.4 13 Cu 324.70 1 - 10 8.3 14 Mg 280.20 5 - 50 16.9 - Pb 283.30 1 - 10 8.1 11 V 318.50 5 - 50 11.5 9 Sn 317.50 1 - 10 10.2 8 - 174 -

TABLE IV Modifications of Experimental Parameters

Conventional Present Photo-electric Photographic method method

Carrier 5Z AgCl(lZGa2O3) 32 AgCl(2ZGa?0_)

SAMPLE/STANDARD 100 tngs 120 mgs

Arc Current 10A 8A

Exposure No delayed exposure 3.0 sees delayed expo- sure programmed in the software for all channels

No early shut-off 3.0 - 7.0 sees early shut-oil depending upon the element.

TABLE V Details of Pre-exposure, Exposure times & the Segment choice.

Element/ Pre-exposure Exposure Exposure start Exposure end Channel (sees) (sees) (segment) (segment)

Cd, Pb 3/0 18.0 2 6

Hg 3.0 25.0 2 8

B, Cr, 3.0 28.0 2 9 Mn, Co, Ga. - 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. "/, •'••: > NET INTLGRATED INTLNSiTY, ANALYTt ANALYIL B IN BLANK

INTERNAL STO. Go IN BLANK ^•W^NET INTEGRATED INTENSITY, INT. GTO. 4100 ----- INTERNAL STO. Go IN STD.S1

1000

3600

3200

2800

2K)0

2C00

1600

1200

800

100

1.0 14.0 21.C 26.0 30.0 TME !S£C(M0S> FIC.1. UOLATILIZATIO*; CURUEC FOR aOROfi 03TAINC0 BY THt OELAYLO EXPOSURE TtCHNJOUC - 178 -

2400 ANALYTE Cr 100.0 ppa IN STO. S1 INTECRATEO INTENSITY, ANALYTE 2200 ANALYTE Cr IN BLANK INTERNAL STO. Co IN BLANK 2000 INTECRATEO INTENSITY/ INT. STO. INTERNAL STO. Co IN STD.S4 1800

1600

5 1400

5 1200

3 iooc

800

600

400

200

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 ANALYTE Mn 10.0 pp« IN STO. SI 2200 INTEGRATED INTENSITY/ ANALYTt ANALYTE Mn IN BLANK

2000 INTERNAL STD. Co IS BLANK NET INTEGRATED INTENSITY/ INT. STO. 1800 INTERNAL STD. Go IN STD.S4

1600

5 1400

5 1200

5 1000

800

600

400

200

0 3.3 1.0 10.9 11.0 11.9 21.0 24.9 28.0 31.9 39.0 TIME (SECONDS* riC.3. VOUTILIZATION CURVES FOR MANGANESE 03TAINE0 BY THE OELAYt'D EXPOSURE TECHNIQUE - 180 -

2100 ANALYTt Cd 1.0 pp* IN STD. 51 INTEGRATED INTENSITY. ANALYTE 2200 ANALYTE Cd IN BLANK

INTERNAL STO. Co IN BLANK 2000 j^S NET INTEGRATED INTENSITY, INT. STO. INTERNAL STO. Co IN ST0.S1 1900

1600

1100

1200

1000

800

600

100

200 1 t 1.0 11.0 21.0 28.0 35.0 TIME (SECONDS! FIC.1. UOUTILIZATION CURVES FOR CADMIUM OBTAINED BY THE DELAYED EXPOSURE TECHNIOUE - 181 -

9 • 7M1 2 3 4 9 « ?a

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 : Hilger's large quartz Wavelength region : 2200-2850 X Diaphragm : A diaphragm having an aperture 3 mm wide used at the collimating lens Slit width : 15 um Analytical gap : A mm 1" Lower electrode (anode) : £ dia u.c.c 7050 graphite electrode to contain 30 mg standard/sample 3 " Upper electrode (cathode) -rr dia u.c.c graphite pointed electrode Excitation source d.c. arc operated at 10 A Exposure time 25 seconds Photographic emulsion Kodak SA-1, 10" X 4" plat* Nicrophotometer Hilger's non-recording microphotom*t*r Data processing Optical densities war* converted to inte- nsities and calibration was mad* using N-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* 1-20 15.9 Mi 2320.0 % Ga 2418.7 X 10-100 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 20 25 30 35 TIHE (SECONDS)-

FIC.1. VOLATILIZATION OF B IN HgF, ; (a) WITHOUT GRAPHITE/ (b) WITH GRAPHITE. - 187 -

0.30 -

10 15 20 25 TIME (SECONDS)-

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 SO 100 -CONCENTRATION (PPM)

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 uranium 1-6 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 equilibrium K2 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 optical density at different conditions, a be

fPAR] X 10.M O.D O.D 530mB 54Onm

0.70 2.0 0 •25D .246 1.13 2.0 0 .430 .410 1.40 2.0 0 ,511 .500 3.50 2.0 0 1.307 1.256 3.50 1.0 0 1.238 1.200 3.50 5.0 0 1.338 1.306 3.50 10.0 0 1.406 1.386 0.52 2.0 3.0 0.100 0.092 0.70 2.0 3.0 0.135 0.125 2.10 2.0 3.0 0.395 0.370 4,20 2.0 3.0 0.785 0.742 0.70 2.0 6.0 0.110 0.100 1.40 2.0 6.0 0.217 0.205 2,80 2.0 6.0 0.429 0.410 4.20 2.0 6.o 0.620 0.600 a) Refera ta concentratlen ef uraniua In the range 1.5 ta 12 «g/l in the aeaauring aalution. b) Refer* ta addition af 0.5 ta 5 al of V.1% solutlan af 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 the reagenta recewnendea cerreapanda ta 0.018yug af uraniua per oa2. ' - 197 - Table II The equilibrium constant calculated under different conditions of reagents in the formation of Uranyl- PAR Complex from U-SDTA. a b c (d) j\j] x 10,K [EDTA] x 10,M [PARj x 10," K

1.4 3.0 1.2 18.3 1.4 3.0 2.0 18.7 1.4 3.0 0.8 18.0 1.A 6.0 2.U 22.0 1.4 1.5 1.0 19.5 1.4 i.O 1.0 22.5 2.8 3.0 1.2 19.5 2.0 3.0 2.0 18.5 2.8 3.0 0.8 1O.0 2.8 6.0 2.0 21.0 3.5 1.5 1.0 21.9 3.5 3.0 1.0 24.0 3.5 3.0 2.5 20,0 3.5 3.0 5.0 22.5 3.5 4.5 2.5 21.6 3.5 4.5 5.0 19.3 3.5 6.0 1.0 24.0 3.5 6.0 2.5 21.6

a) Variation of uranlua in solution is 4 mg te 10 mg par litre. 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.

^ Complexing PAR,0.1% Uranium found • solution, ml in 25 ml. ml per 25 ml Bn/litre a b 1 1.0 1.0 3.20 3.40

* 2 2.0 1.0 3.14 3.18 CO. t> 1.0 0.6 3.06 3.12 u 1.0 0.4 3.01 3.08 5 1.0 1.0 0.272 0.275 6 2.0 1.0 0.26y 0.267 7 1.0 0.6 0.270 0.274 8 2.0 1.0 0.197 0.203 9 2.0 1.0 0.205 0.212

1) a and b reler to two dilutions or tho original sample used for the measurements. For SI.No, 1 ted, tho dilutions are 1250 and 325 respectively and for 5 to 7 tho dilution are 100 and 50. 2) The sample of serial numbers 1 to U was analysed to contain 12.5 ga/lltre of iron* 3) SI.No. 8 and 9 refer to a synthetic mixture of uranium ana 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 of couplexing agent and PAR and the values agreed within 0.5%. - 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. M Kicr*. 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 matrix interference 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 on analysis of Pu by ICP-MS* But I have seen that ICP-MS has been already adopted tc glove-box operation in the Institute for Transuranium Elements (European Atomic Energy Commission) at Xarisruhe, Westt .Germany. They were analysing Am in the active waste^solutions. . *

S.M. Marathe t What is the limit set for maximum solute concentra- tion? Do you experience clogging of nebullser?

T, 3, .Mahalingam si ihlnk that you are referring to the maximum 3ample or aalt concentration. Generally a 0.1* solution is easily handled. "Matrix interferences are more at higher concentrations.

We did not expirienoe any clogging problem even with 1* solution of sodium nitrate.* Published literature indicates that even with refractories, unwrnxxcixm no clogging problem was observed, when the sample concentration was kept at 0.1* or leas.

K. Syamaundar i What la the aample sise of uranium taken for the rare earth determination? What la the. detection limit for Gd fen achieved with that sample slse? What la the throughput of samples for analysis? - 202 -

T.R. MAHALINGAM : The detection linit for Gd is 1.8 ppm. But once the uranium is removed by solvent extraction, the detection limit has been found to improve to 0.002 ppm. About 5 samples could be analysed for ten elements in about one hour.

Paper Ho« 3

N. MAHADEVAN ; Why do you want a better or a superior spectrophotometer for simultaneous analysis of TJVT and and UIV in your PIA system?

A.H. PARANJAPE : If both UIV and UVI are to be measured from single injection of sample then simultaneous measurement of TV VT U at 65Onm and U at 420 ntn would be required. The detector we have used is for absorbance measurement at single wave length* Simultaneous measurement would require a stopped flow technique combined with a spectrometer capable of automated scanning at two wave length. S.K. AGARWAL t Can the flow Injection analysis technique be used IV VI for determining the per centage of U and U ?

A.H. PARANJAPE i Y«e, TJVI does not interfere at 650nm where UIV is measured. So it can be analysed without difficulty. If VT • IV U is to be measured in the presence of U then correlation of absorbance at 410 nm for the presence of UI V is required at nee it Interferes at this wavelength* - 203 -

Paper No. 4

H.C. JAIN i What is the lowest limit of B and Cd in the high purity uranium which is used for making master?

S.S. BISWAS t High purity U-Og used in preparing the MASTER STANDARD for B, Cd etc. is examined by optical-emission spectrograpbic method. If the spectrum does not show lines due to B and Cd, under normal exposure condition, it is inferred that these elements must be less than 0.05 ppm. AbOTe this level, B and Cd lines will be seen in the spectrum.

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 final callibration graph* for B and Cd using certified batch of U~0g, will further confirm that both these elements must be present at 0.05 ppm level. However, we would prefer to check these values with S3-MS method.

A.B. PATWAHDHA? » If blank contains impurlt. nil this be subtracted?

S.S. BISWAS t Yes, the "blank subtraction" corrects for the "electrode blank" also i.e. for example Mg. Further, it corrects for the residuals In the matrix and the noise contributions from electronics circuitry. The overall effeot can be seen by the it R3D In the table given in, the text* S»SSIOH III A

MINING AND ORE BENE7ECIATION

Chairman : Shri A.O. SARA SWAT Director AMD Reporteur* Dr. V.N. Pandey UCIL - 204 -

DEVELOPMENT OF niNINn A.T JMQUGUQA

By

J. L. Bhasin Chairman & Managing Director Uranium Corporation of India Ltd 3adugude

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

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

* 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

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.

-HHBL HUHOTltt l*.'.'.*l ..aeblat X7*nlt« aobltt Ortbody 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

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 WATER SPRAYING

. 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

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 -

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 rilteratlon 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 /

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

FlC, II - STOPPIN6 SCHCUATIC PIAGBAM OF PROPOSED H — RfdULATOR VT WtLTWOBK OF jAT»Haiit>A V - TJOOR TO - TOP PIMVl - 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

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 '"

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 frc 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<

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 satisfactorily* Figure—10 shows the performance of the mine during the last S years. It may be seen that during these years, the performance is above 85flC, which is quite achievement for an underground nine.

200,000 2.C8.7/9 095 7.

2.40.0O0

z.oo.000

U 1.60.000 Q O 1,2 0.000

8 0.000

+0.000

'900-09

FlG.10. PERFORMANCE OF JADUGUPA 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 70-80 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 % Rock quality Less than 25 % Very poor 25 - 50 % poor 50 - 75 % Fair 75 - 90 % Good 90 - 100 % 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 230 and E2 "!)• 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 ^and 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 SOM 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

LEAO BODf

CABLE

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

LEAD SHJELO PROBE SCINTILLATION PROBE *%3 - 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 erroneous surveys 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 -

oL mechanical 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 CONCENTRATE FROM COPPER TAILINGS.

• •• *•» S.CHAKRABOPTY, U.K.TEWARI & K.K.BERI

Association of uranium mineral with copper ore of Singhbhum Thrust Belt was known for quite long time and efforts were made to recover them economically from time to time.

The first attempt to recover uranium from the copper tailings was made in the Moubhandar works of M/s HCL/ICC ( the then *Indian Copper Corporation* around the fifties and sixties. The experiments met with little success mainly because of the data available in this field was very meagre* the shaking tables used were of primitive design with very poor efficiency* Jigging, flotation, tabling etc. were also tried, but in vain. The project was subsequently abandoned. With the opening of the Surda mines in the early sixties, which reported higher uranium 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 copper tailings of the Moubhandar works of M/s HCL/ICC was

* Deputy.Supdt(Chem) •• Addl,Supdt(Chem) ••• Chief Mill Superintendent Uranium Corpn of India Ltd P.Os Jaduguda Mines Singhbhum, 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 transparent minerals s 0,4% Other apaque oxides 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 mineral from copper tailings will be quite significant for national requirement*

EXPERIENCES IN THE URANIUM RECOVERY PLANTS & THE HIGHLIGHTS THEREIN X

Use of High Density Polyethylene Pipes :

For the first time in this company, " high-density polyethylene pipes'were used in the slurry lines in place of the conventional rubber-lined pipes at the Surda Uranium Recovery Plant. Not only the cost of the H.D.P.E pipas wara 19*9, but they were also easier to install easier to maintain & have a very long life* Initially these 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 a result, with tha fluctuation in temperatures, the pipas wara getting cracked* These pipas wara then laid loosely on mild-steel trays wifth loose clamping. This gave the necessary room for expansion of the pipas caused by tha temperature 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 pumps ara 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-bank to 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 RECOVERY I 6H*VPUIP P/STfilBurOR r 2 4*34: °K PLANT BEFORE MODERNISAT ION J. b_S} 4- 5 T/jfi'/NQS S PJTf 11- - •—-e- a— n w \ V 7 \ -41 e B -< \ / / 1 4 A -\ 4 H B Q

/ 7 \ Y 7 / \ T -1h B B ^ / 1 A \ 4 A \ 4 1 (TJ H tea

1 / 1 7 \ / \ V CC •1 B 9 -t *- =) of \ / A C O I D 4 \ 4 ^ L B -1 a • I 0 a a Q J & •«

/ / r \ 7 \ LAYOU T A B R -I / f A \ 4 A \ 4 — -IB - TABL E

• 7 / \ Y 7 / \ -< K "V B B h / \ A \ 1 4 •i 1- B a —1-

l 1 r / \ 7 / \ B fi A \ / A A \ I ^^ —B -B-

- 275 -

/*« S FLOW SHEET FOR C.TS/BM&/CBC FLOW SHEET FOR CTS/BMS/CBC SYSTEM TESTING CIRCUIT NO- * SYSTEM TFSTING CIRCUIT NO- 1 COfPfR TAII INGS

rarrtK TAILINGS I TAB'LE I "»IE TAt5i

I TABLE COWCl ICTS SCREEwi COARSCREJECT IC.TS. SCREEN! MO. »i r I CT.S- I C T i f INES | ICT.SCOAHSEI

1 1M SEPARATOR! It M SEPARATOR^ —IH TAILS REJECT I (ROUGHER CONO COMC- IIIOUS ER CONCl Ml' DLINO 1 __HJPPUf« 1 D I {QMC

I HHAL CONCI IflWAL COHC. I

•'"I MATEBIAL BALAWCt t\. Ow SHEET — i MATERIAL BALANCE fQ» 'lO» SHEET -I

I'M.* I lOtPOO'lWOI

CIS. KRCtNl »O t*B I En CT» FINES icT.i cotim |4144>0«>*l«

| TttLE >—- tun >—| mm • » M»I««TO«| K»3

COMClHTHtTt • •OUCMCR COHC |HI4|OBH'0 I ll-IOIO^li III'97 ™^Jr^ mni>t ' I 1 C.

|H0|0-CHp4

•I coxet N r«4if • 'I t u,o, ••iijr »» 1 0-9411Jl-H I 'mil. CONC r>- 1004 - 276 - TAB IE -'w'

: "Metallurcical data on RDCC and Tables

o Tonnes Grade % % Distribution Kg U3O8 O t 8 g « dry ore U3O8 of U3O8 at the at the • stnee staee

1. Feed 1to Hydro cyclone 1000.0 0.0104 100.0 104.00 2. Hydro cyclone over - 345.0 0.0097 32.7 33.90 flow 3. Hydroclone underflow 655.0 0.0107 67.3 70.10 4. IIDCC <:onc. 01 166.7 0.0197 31.2 32.SO

5. RDCC ( :onc. 0? 83.3 0.0169 13.5 14.10 6. RDCC ''ails 405.0 0.0062 ?4.0 23.20 7. 8 Wet Tables 8.2 0.3055 24.1 25.05 Cone. TC 1 s. 8 Wot Tables 158.5 0.0052 8»0 8.20 Tails IT 1 9. 4 Wet Tables 4.1 0.2347 9.3 9.62 Cone. TC 2

10i.4 Wet Tables 79.2 0.0052 4.0 4.10 Tails TT 2

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

frpn In- '•:: «-.trli)uti in xhe various size trrcx ions of the feed the t,~!>le the c -ncontr^te ,-rp n!von bojiw, ilAWAl iiae ..• C -> Nc r. N "a A T £ ivt. *"- •.t. ^, U308 •. Wt. r. ..t. % U?O8 "& U3O8 In - in Uistri f rrc , frrcti-m button

+4* 1 .0 1. 0 0. CC1S 1.5 1.5 0.01T5 0.C5 -18 -»-65 3 .0 4. 0 0. 003? i.i 1.5 3.0 C.0286 0.57 .- 65+1C0 ? .5 13. 5 0. 0029 3.1 ?.5 5.5 0.0210 0.7P -IOCK-O 1.".5 27. C 0. "03? 5.0 6.0 11.5 0.02S8 ?.3O -iscrroo 14 .C I-1. 0 0. 00" ? 5.? 10.5 •'.2.0 C.':292 4.OS

-20Z+2.5 ro.5 ol. •5 w • 0051 11.0 30.5 C-2.5 C.030'.' 1.-.17 -?:S 3S .5 100. 0 c. 01,7 T... 4 •i7.5 1CC.C 0.15«9 79.93

.-'ceo ri-rtie : •3. 009 6 Cane 0.074C

" ••p.l.;c5 in in* i Lider frrcti n» of« higher; tho vBluas in *.h« -3?5 ons ,-ro almst rio ^lo *• i \*\a ^rwes of *.he food or Vhe concent>rtt.

,. ..: SAT :•. . '-t • -..t '". L'.'-O? V.t '• ',it '.. V3JP '.. U C8 in -intri- in L>i»tri. fr?ott^n '-•itti'.n fvrctl-v. ' t:ti-"i T50 l.C 1.0 l.'l? 15 .5 IP.5 0.C161 4.01

•7 ••") ' . 0 •i.o 0.0061 1.95 .0 IS. 5 0.0.314 1.51 *1CC l'.O 0.0-38 S.51 11 .0 T9.5 0.0?18 5.02 t-150 :'-\5 :'4.5 0.0047 9.46 17 .5 47.0 0.0??»0 e.ie T'JC 15.0 ?i.5 O.OOSo 12.10 17 .0 (54.0 0.0385 10.5? 17.5 57.0 0.0050 14.10 15 .5 71.5 0.0673 16.77

-3.-5 n.o 100.0 0.0090 55.40 PO 5 100.0 53.30 • 0.10P0

'.'.COS? Tone. Or ado 0.06.?2 - 278 -

TABLE - III

Comparative fractional recovery of Uranium at R.U.R.P and M.U.R.P.

Fractional recovery in %

Size RURP MURP

+50 52.08 85.11 +70 22.34 23.16 +100 9.50 30.69 +150 20.00 25.91 +200 33.18 26.18 +325 43.76 35.77 -325 46.76 28.96 - 279 -

TABUE

tg g/toE TESTPC (BIS

3. 0 F.td 0 Co«™ 0 Finn B»S « «t. t ft DM cone. 0 DHS T*11.4 B>St»ce. C 1H hwttn Ho. 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 «.IJ * MU. 2.9* 300

2. 0.0074 0.00M 0.0090 *0 0.0231 C.O06* 49.lt 5 ItaU. 2.5» 300 0.0044 «

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 Sl.« 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 0.0103 Q 6. C.0C73 e.ccm c.ocei* so c.ci2o o.co6i es.w 4 »»it«. a.s 300 c.0094

Th> at. ff-*etlon of film ind eoan* w»» fo-jnd to to SOX «rriox.

TABU N.. .

1ST CTAg TE T1W0 IKftWAiB.

a.lto. F*t* to t*b|t* Cenetntnto Tall* %

I. 0.00a 0.04 0.004* 30.37

2. 0.0OS*. 0.CB04 O.OOSO 19.3

3. 0.006 0.044 0.0C49 22.49

4. 0.0067 o.o«a 0.0O7 17.0

s. 0.0064 0.03 o.oas 16.6*

&. O.OOS 0.030 0.0C4 30.00 - 280 -

TABU No. r.yl

1ST STAff TESTING (CBC FTfFOHWICE)

c cone. I coc T.m.« oc ccc rtcov.nr 0 a.*. « coc *"* 6 Slop. 0 OrbUtl B*lt J 1"3 Cone, X « * I5 * % 6 «p»»d

300 1.86 1. 0.(266 0.0487 O.OOeB B.eyeltd to 86.5 fMd

B0.3 300 1.86 ,, 3. 0.C231 0.O04 0.0372

8 3. 0.0237 o'.OA8 0.0059 86.1 300 I- * •»

4. 0.026* 0.0B2 0.00B9 77.» 320 X.96 „

0.0273 ,0.0504 0.0058 08.9 300 i.eo ,,

0.C120 0.0180 C.006 -do- 75 1.0 260 1.9 ,,

TASLi IIP. VII 2ID STAGE r-ffirg. nms

S.Hs. Ttbl» erne. n. % »f TWbU TtlU T«bl« itcaviiy 1 UXM eone.ntntt JH!3S2__ C

I. e.cces* 0.CM7 4.C5 0.0066' 23.11 0.0071 0.OCT5

I. coon* 0.07* 3.16 0.0068 25.es • C.OOBf .

3. C.0105 0.0869 3.66 0.0083 32.06

4. 0.0098 0.0496 4.81 0.0074 24.36

5. 0.007 0.06 2.58 0.0051 '22.03

6. C.CW6 0.O5J7 5.17 0,O>K> 25,3*

ErtUtttd - 281 -

TADU .'u STAGS r 3TltC ( PERFOflWiCE OF BM3)

Sl.t CTS coars»s •it.it fln»t 5Hi conc.l 2:iS tall 1BMS Hccot W IVsT cr: fint7T (Tab.tail * U3J8 1 S U338 ;ycl«t H X U338 basis) I * rtin*! Slontl i. 0. 0066' 0. 00<9 0.C075 67.08 0.0134 0.006' 33 4 2 300 0. 0071 0.0039

2. 0. 0068 0. 005* 0.0081 51.97 0.0120 0.0068 34 4 1.9 300

3. 0. 0083 0. 0069 0.0087 55.09 0.0137 C.0072 38.96 4 1.9 300

4. 0. 0074 0 0054 0.0097 55.08 0.0145 0.0083* 37.36 4 2 320 0.0071 0.0071 5. 0. 0051 0 0047 0.0067 50.0 0.0110 O.C044 55.82 3.3 1.29 300

6. 0 008 0 .0071 0.0098 53.52 0.0151 C0066 44.11 4 1.5 300 0.0074

• Eitiaattd

TABLE JO.'

T-CTHO P.'.A:>:£ OH CSC)

ST. --- ••« i. Cone. <-3S Itil CiJ -itari- •*arifa>icr» No. ',i-'-'S cane i. ••3?.:\ * U336 • lop* ursiial a*lt

1. 3.9134 0.32* 0 .0357 ilecycl** \o 74.0 2 130 1.34 n/nt. 0 .0233 # 2. 0.OI2C 0 ,0351 0 .0058 69.7 2 300 1.83 •/at.

m 3. 0.0137 0 ,0207 0 .0072 77.36 2 300 1.93 •/mt.

4. 0.5115 0 .0213 0 .0074* • 73.6 . 2 300 t .83 n/nt. 0 .0C53

5. 0.0110 0 .023 0 .0061 • 43.99 1.9 320 1.39 a/at

6. 0.0151 0 .029 ' 0 .0091 • 61.73 1.5 320 1.39 •/nt

• Eatlaatrt - 282 -

TABUE NO.' X

1ST Stao* Ttttlna (Ovrall

51. rctd KCC. Lonbincd ola H«c. Overa11 reeovrv «• % No. X U308 rtcovary (coara* XT fin* At coax** Total (en flno» baala) (tag* • tag* )

1. 0.0069 42.13 86.9 36.44 23.39 20.06 10.908 30.97

2. 0.0074 49.86 80.3 40.03 19.30 24.34 9.99 30.33

3. . 0.0084 47.12 86.1 40.97 22.49 23.83 8.28 32.16

4. 0.0083 91.92 77.69 40.31 17.0 24.04 6.86 30.90

9. 0.0091 90.24 88.90 44.66 16.68 =-.99 9.66 34.81 * 6. 0.0073 99.69 7S.0 41.76 20.0 23.98 8.90 32.48

TABIE Wo.

2io nxr Trarc levr MI tOFO v*wa)

S. tiH tt^r»U CDCr*c. CrUnrtric. TaMt»«. Owtmll wcawtv -1 Ik. JHO08 * « (•» «!»• fc*tl») K MliM labllab * K •tap itw.

1. O.CCCS* 33 2S.it 13.69 25.11 ».O 0.007}

0.C0M* 34 ».7 M.ca C.« •.CO*

3. 0.C1OJ -.96 77.36 3C.14 M.0» I3.» 22.06 K.49

4. o.eow 37.36 70.6 34.36 13.31 34.M 37.97

9. 0.CC7 55.19 tr.fi X.«3 ».O3 22,» 22.0S 44.U APPENDIX 1

The following tests were conducted on laboratory scale on Mosaboni copper tails: 1. Low Acid Leach Test The tailings from Mosabani concentrator (feed to MURP) were collected from time to time in batches (about 10 Kgs) and sent to the laboratory for leaching tests. The wet samples were filtered, dried and mixed thoroughly. 10 Kg sample (in two batches) was taken for each test. Pulp density : 60* solids pH i during 1st hr 2.3 after 7 hrs 2.1 - 2.2 Emf 450-500mV. a) Using ovroluslte for oxidation

Sample No. Temp:C Acid Pyrolusite Head Leach Leaching and date (Ambient) consumption consumed Assay Residue Efficiency Kg/T %u3o8 1.(25. 3. 88) 29 17 .1 5.20 0.0074 0.00195 73.6 2.(30. 3. 88) 31 15 .7 5.65 0.0102 0.0024 76.5 3.(06. 4. 88) 36 14 .3 5.00 0.0071 0.0015 78.9 4.(20. 4. 88) 30 15 .3 5.25 0.00996 0.0023 76.9 5.(29. 4. 88) 32 10 .0 4.00 0.0082 0.0027 67.1 (pH 2.4 for 8 hrs) 6.(06. 5. 88) 38 13 .7 2.60 0.0Q92 0.0017 61.5 b) Usina Ferric Sulohate Lea, china Head Assay t 0.0092% Pulp Density : 60% solids Temp. : 36°C Emf t -450 mV

H2S04 consumed t 10.4 Kg/T

Fe2(SO4)3 consumed t 6.5 Kg/T Leach Residue t 0.0014% U Leaching 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 CENTRE HYDERABAD

Petrology of the ore plays a key role and Influences the technology and economics of the processing of the ores. In the absence of high grade uranium resources, low grade ores constitute a significant uranium resource in India, and preconcentration before leaching is necessary in rendering-these resources viable, and in minimizing environmental deterioration. Copper tailings of Singhbhum Shear Zone are such lean resources and UCIL is exploiting these ores by setting up preconcentration plants at Rakha, Surda and Mosaboni. Intensive petrological work has been carried out on these ores. Nature and distribution of uranium minerals is studied by microscopic examination of thin and polished sections: and the mineralogical composition and distribution of uranium values in various size fractions have been determined by a combination of sieving, heavy media separation, radiometric assay and microscopic examination. In the light of the petrological data, various problems involved in the preconcentration and leaching of these lean ores, and different technical options 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

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 ranium Deposits, Proceed. Symp. IAEA., Vienna, pp 625-635. Banerji A.K., 1969, A Reinterpretation of Geological History 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 on Beneficiation of Rakha Mines Copper Ore (for NMDC), BARC/ Met / 10.

Ghosh A.K. and Banerjee A.K., 1970, On the Nature of Petrogenesis of Dhanjori Lava near Rakha Mines, Singhbhum, Bihar. J .Geol. Soc. India, v 11, pp 77-81. IAEA., Vienna, 1980, Significance of Mineralogy in the Development 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 Rao N.K., 1987, Recovery of Uranium values from Copper Tailings of Mosaboni • Upgradation of BMS Concentrate on Cross Belt Concentrator. A report submitted to UCIL. Jha R.S., NataraJan R., Sreenivas T., Sridhar U. and Rao N.K., 1988, Amenability of Uranium Ores of Singhbhum to Wet High Intensity Magnetic Separation. BARC/ 1-947. Rao N.K., 1977, Mineralogy, Petrology and Geochemistry of Uranium Prospects from Singhbhum' She&r Zone. Bihar. Ph.D.Thesis, Banares Hindu University. , Rao N.K. and Rao G.V.U., 1983a, Uranium Mineralization in Singhbhum Shear Zone, Bihar. I - Ore Mineralogy and Petrography. J. Geol. Soc. India., v 24, pp 437-453. ; - 296 -

Rao N.K. and Rao G.V.U., 1983b, Uranium Mineralization in 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 Mineralization in Singhbhum Shear Zone, Bihar. • IV - Origin and Geological Time Frame., J. Geol. Soc. India., v 24, pp 615-627.

Singh H., Padmanabhan N.P.H., Rao N.K., Sridhar U. and Rao G.V.U., 1981. Differential Comminution and its Applications 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 Engineering Analysis of Uranium Recovery from Copper Tailings by Wet Tabling, 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 a Gravity Concentration Process for Improving Uranium recovery from Copper Tailings, BARC/I-853.

Sarkar S.N., 1980, Precambrian Stratigraphy and Geochronology of Peninsular India : A review, Indian J. Earth Scl., v 7, pp 12-26. - 297 -

Table I. Mineralogical Composition of Typical Feed (Wt%) Mineral % SURDA ORE RAKHA ORE MOSABONI ORE Quartz 62.3 69.2 51.8 Chlorite 22.3 9.9 39.2 Apatite 2.3 2.6 1.7 Tourmaline 3.6 6.4 0.8 Opa-f Magnetite 3.2 8.5 \ ques\ Sulphides 5.8 3.3 J ' Others 0.5 0.1 0.9

Table II Petrological Data on Rakha Copper Tailings +140 -140+270 -270 Total (>104pm) (>52pm) (<52pm)

WeightX 45.1 32.9 22.0 100.0 FEED *U3°8 41 71 220 90 Distn.X 20.5 25.9 53.6 100.0 BRL Weight* 35.7 22.67 13.13 71.5 (Quartz) - 2 63 (2)* Distn.X 0.5 9.16 9.66 (0.3) WeightX 8.8 8.06 7.19 24.05 MIL 91 62 96 (Micas* 3 (62) Apatite) Distn.X 8.87 5.54 7.65 22.06 (4.9)

uru WeightX 0.61 2.17 1.67 4.45 Win Distn.X 11.63 19.86 36.79 68.28 Figures in parenthesis indicate probable unliberated values - 298 -

Tablelll.Petrologlcal Data on Surda Copper Tailings.

+ 100 -100+270 -270 Total

Weight* 34.3 44.8 20.9 100.0 FEED 67 104 276 127 Distn.X 18.1 36.6 45.3 100.0 BRL Weight* 35.9 27.37 10.35 63.62 (Quartz) eU3°8 24 2 43 (2)* Distn.X 4.9 0.43 3.49 8.82 (0.2) Weight* 7.68 14.65 9.28 31.61 "U3°8 170 67 120 (67) Apatite) Distn.X 10.28 7.7. 8.74 26.73 (4.9) WeightX 0.72 2.78 1.27 4.77 MIH Distn.X- 2.92 28.46 33.06 64.44 Figures in parenthesis indicate probable unliberated values. - 299 -

Table IV. Petrological Data on Mosaboni Copper Tailings.

+ 100 -100+270 -270 Total 0147pm) (>52pm) (<52pm) Weight* 40.2 41.2 18.6 100.0 FEED eU3°8 86 102 190 112 Distn.* 30.9 37.5 31.6 100.0 BRL Weight* 27.66 21.34 6.85 55.85 (Quartz) eU3°8 23 5 27 (5)* Distn.* 5.69 0.95 1.65 8.29 (0.3)

urr Weight* 11.5 16.65 10.30 38.45 nXL eU3°8 162 96 129 J o (Micas* (96) Apatite) Distn.* 16.65 14.24 11.89 42.86 (8.8)

MTU Weight* 1.04 3.22 1.45 5.7 n In Distn.* 8.57 22.29 16.06 48.92

Figures In parenthesis indicate probable unliberated values. - 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 Section Bhabha Atomic Research Centre.

The existing practice of preconcentration of uranium values by wet shaking tables offers limited scope for improving their recover/ from coppqr plant tailings, particularly at Mosabani Uranium Recovery Plant (MURP). The overall recovery at MURP is only 18-22%. Extensive studies on improving the recovery of these values using fine gravity machines have been carried out in the Ore Dressing Section laboratory and an integrated gravity flowsheet arrived at. A pilot plant using full scale machines was set up at MURP with the help of UCIL engineers to test the suggested flowsheet and collect data for the scale up and design factors.

The feed was classified into fines and coarse sizes using C.T.S 268 screens and the fines containing higher distribution of uranium values was processed on the Bartles Mozley Separator (BMS) and the Cross Belt Concentrator (CBC), while the coarse fraction was treated on conventional wet shaking tables, supported by matching conditioners and pumps.

The findings of the laboratory studies could not be directly scaled up at the pilot plant stage due to dissimilarities in the area of B.M.S and variation in feed characterstics, thus necessitating certain changes in the operating parameters of B.M.S. and further optimisation studies of the same for maximising the recovery of uranium values.

The pilot plant studies have shown that an overall recovery of 35-40% is feasible. This does not Include the additional recovery obtainable by recovering ultrafine 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 analysis of Uranium Recovery from copper tailings by Wet Tabling". 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. "Laboratory investigations on improving Uranium recovery from copper tailings by gravity beneficiation". Investigation Report 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 improved gravity concentration process incorporating Bartles Machines- Bartles Mozley Separator and Bartles Cross Belt 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 Mosaboni copper tailings". Investigation Report submitted to UCIL, 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 of Uranium 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 of Mosaboni: Upgradation of BMS concentrate on Cross Belt Concentrator". 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 fine particles". Paper presented at the Seminar on Recent Developments in Mineral Engineering. April 1989 Jamshedpur. - 31? -

Table I. Uranium in Copper Tailings.

Surda Rakha Mosaboni Total I.Tailings at present Throughput TPD. iOOC 2700 4700 Million Tonnes/year. 0.3 0.3 0.8 1.4 2.Tenor U_O_ppm 0 0 Range 90-120 90-110 70-100 Average 105 100 85 3.Contained U,Og in ions. At min. grade TPY 27 27 57 111 At average grade TPY 32 30 60 131 4.Contained U-Ogln tonnes Per million tonnes of ore processed. At minimum grade 90 90 70 At average grade 105 100 85

Table II. Distribution of U Values in different sizes in Mosaboni tailings and and applicability of gravity beneflciatlon equipment.

Particle size range (MDI) >100 100-37 37-7 <7

Mosaboni % weight 16 26 48 10 Wet shaking table Yes Yes No No B.M.Separator No Yes Yes No Cross Belt Concentrator No Yes Yes No Hydrocyclone No No Yes Yes Table III. Laboratory Test Results Using Integrated Gravity Flowsheet. Lab. test code MSB-PP MPP-1 MPP-21 MPP-22 Feed assay 100 95 85 70

BMS Wt.% 17.6 19.0 17.5 17.2 RGHR assay 340 222 190 190 Cone. Dist% 59.8 44.5 39.3 40.7 BMS Wt.% 7.8 7.6 5.38 5.0 CLNR assay 690 470 528 410 Cone. Dlst% 53.8 37.6 33.4 29.3 Wt.X 2.77 0.63 2.06 1.3 DT 367 610 410 Cone. assay 530 Dist% 10.2 4.0 9.92 10.2 MIXED Wt.% 10.57 8.23 7.44 6.3 • TABLE assay 605 480 495 434 Cone. Dlst% 64.0 41.6 43.3 39.1 CYCLONE Wt.X 2.5 2.73 2.75 3.0 OVER- assay 130 210 200 200 FLOW DlstX 3.3 6.03 6.49 8.6 Wt.% 13.1 11.0 10.2 9.3 U.Cone. assay 515 413 415 360 DistX 67.3 47.6 49.8 47.6

Table IV. Equipment Scale-up Criterion. SI. Equipment Units Lab Test BGML PLANT No. TESTS DESIGN 1 CTS Screen TPH/m - box capacity 2.55 2.3 B.M.Separator 2 n»3/hr/m2 0.34 capacity 0.35 0.35 Pulp density Xsollds 8.0 11.0 10.0 3 Cross belt 3 2 - concentrator m /hr/m 500 Kg 500 Kg 4 Shaking table TPH/nT 0.13 - 0.15 5 ' Hydrocyclone Size mm 75.0 100.0 150.0 Pressure Kpa 100.0 150.0 340.0 Capacity m3/h 4.0 10.0 20.0 Type — Dorrci one Rletma i Table V. Classification and Uranium Distribution in CT5 streams.

Feed Fines Coarse Wt% in Dist.in Code- Nn f 1 np<* f 1 r>A

Table VI B.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. )

Code Flow Feed Concentrate Recovery No. RPM rate U (lpm) TPH 3°8 TPH WT% U3°8 Stage Overall 2/3 225 428 1.4 109 0.34 24.0 254 55.2 34.5 4A/3 225 428 1.9 100 0.37 19.4 254 49.3 30.3 4B/3 240 420 1.9 100 0.43 22.4 256 57.0 34.5 6A/3 225 435 2.5 96 0.34 13.4 335 47.0 36.1 7A/3 225 420 2.27 10B 0.48 21.1 325 63.0 41.6 8A/3 225 430 2.3 84 0.52 22.6 203 54.2 31.6 9B/3 225 460 2.02 77 0.37 18.0 181 42.3 29.3

Table VII. Levels of C.B.C. Variables. Upper Lower Belt speed M/mln. 1.52 1.39 Percent 30 25 Shear rate rpm. 240 225 Feed slurry flow rate lpm 60 45 - 315 -

Table VIII. Results of C.B .C. Tests (All assays are by chemical analysis in ppm) Feed Concentrate SI. No. U3°8 X weight U3°8 X Dist. E.R. *1 147 24.1 362 59.4 2.46 2 159 22.5 414 58.6 2.60 3 156 39.4 283 71.8 1.82 4 150 14.0 545 50.8 3.63 5 140 33.8 279 67.4 2.00 6 140 21.3 420 63.9 3.00 7 146 51.0 220 77.1 1.51 6 140 52.2 212 79.0 1.51

Table IX. Results Obtained on Modified Test Facility at MURP (carried out by UCIL) (All assays are in ppm) Experiment No. 1 2 3 4 F Wt.X 50.0 50.0 57.9 50.0 FINES assay 76 90 100 99 S DlstX 55.0 60.8 68.9 59.6 E Wt.X 50.0 50.0 42.1 50.0 COARSE assay 62 58 62 67 D DistX 45.0 39.2 31.1 40.4 Wt.X 6.0 11.9 11.5 9.7 BMS Cone. assay 266 231 237 266 DistX 23.1 37.1 32.4 30.9

Wt.X 2.8 4.4 5.13 3.22 CBC Cone. assay 487 504 458X 622 DistX 20.0 29.8 28.0 24.1 Wt.X 2.0 0.88 1.5 1.25 Table Cone. assay 400 504 440 458 DlstX 11.6 6.0 7.94 6.9 Wt.% 4.8 5.28 6.63 4.47 Combined assay 400 504 454 576 Concentrate DlstX 31.5 35.8 35.9 31.0 - 316 -

Table X. Liberation Analysis of Mosaboni Feed. BFL MIL MIH

Wt.% eU3O8 %Dist Wt.% eU3°8 %Dist Wt.% eU3°8 %Dist

LAB SI 63.17 20 13.46 34. 10 96 35.41 2.73 916 51.13 OST SI 61.60 47 30.80 35.10 117 43.70 3.30 727 25.50 OST S2 50.90 43 23.40 43.70 104 48.60 5.40 486 28.00 OST S3 62.00 45 33.20 35.90 98 41.90 2.10 996 24.90 OST S4 54.80 37 20.20 41.00 117 47.00 3.40 982 32.80

Fig.l. Integrated Gravity Flowsheet for Uranium Recovery from Copper Tailings. Feeid |CTS WET SCREENS|

i 1 i Fines Coarse I 1 BARTLES MOZLEY WET SHAKING SEPARATOR TABLE WILFLEY or DIESTER

Rougher Tailings DT Cone. DT Tails Concentrate

BARTLES CROSS BELT -Tailings- CONCENTRATOR

Middlings Cleaner HYDROCYCOLNE |- Concentrate -Sand-

Sl lines

Mixed Table Cone. 1 Final Final Waste Uranium Cone. Tails - 317 -

Fig. 2. M0SABAN1 PILOT PLANT SET UP

BUS FEED BOX

PULP OST

WILFLEY Cu. TAILINGS TABLE

TAILS -REJECTS PUMP CLASSIFICATION B.M.S. STAGE COARSE TABLING SAMPLINO 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 uranium minerals, pre-concentration of copper plant tailings of Singhbhum area have been investigated in a pilot plant scale wet high intensity magnetic separator (WHIMS). The experiments were aimed at maximising the recovery of uranium values in the magnetic fraction, as it would greatly reduce the quantity of the material to be processed by leaching and improve its grade to higher economic levels. The variables studied include magnetic field intensity, matrix drum speed, feed slurry flow rate and its pulp density. The results of these investigations have shown that 75-85% of the contained uranium values could be recovered in 45-55% weight in the magnetic fraction in the case of copper plant tailings from Rakha, Surda and Mosabani. The losses in the non-magnetics were primarily due to the ultrafine liberated uraninite particles not collected by WHIMS due to machine limitations and the values occurring as fine inclusions in quartz.

Improved recovery can be obtained by offering higher field gradients and preventing loss of very fine liberated uranium values. High gradient magnetic separator (HGMS) offers higher field gradients. A test in HGMS has indicated superior results in comparison to WHIMS.

. 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 probability of 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 of Uranium recovery from Copper Plant Tailings by wet tabling" ,BARC/I-771,1983. (2) Jha, R.S., Singh, H., Natarajan, R., Rambau, Ch.,and Rao, G.V.O., "Investigation on Gravty beneficiation of Uranium Fines", Int. Conf. on Recent Dev. in Met. Res; Fund. 4 App. aspects, IIT Kanpur, Feb. 1985, Proc. Vol PP 23-29. (3) Corran, I.J.,"A Development in the Application of Wet High Intensity Magnetic Separator", In " Fine Particle Processing" vol II (Ed. P.Somasunderan) AIME, 1980, pp 1294-1309. (4) Nesset, J.E and Finch, J.A, "Loading equation for High Gradient Magnetic Separator and Application in Identifying the fine size limit recovery", ibid., pp 1217-1241. (5) Zimmels. Y., Lin, I.J., and Yaniv, I.," Advances in Application of Magnetic and Electric Techniques for Separation of fine particles", ibid., pp 1155-1177. - 328 -

(6) Gerber, Richard., and Bris3, Robert R.. "High Gradient Magnetic separator", Research studies press, John Wileyfc Sons Ltd,1983. (7) Jain. S.K.."Ore Processing". Oxford IBH Publishing Co Pvt Ltd. New Delhi. 1986. p 352.

Table I RESULTS WITH MURP SAMPLES

Code MAG+MID NMAG Operating Parameters Xwt U %R ER %wt XR (T.rpm.Xsd,lpm) 3°8 °3°8 ppm ppm WH/2 62.9 150 86.3 1.4 37.1 40 13.7 1.8. 4. 30. 13 WH/3 53.1 180 87.1 1.6 47.0 30 12.9 1.8. 2, 10. 12 WH/5 51.1 158 80.5 1.6 48.9 40 19.5 1.5. 2. 20. 14 WH/7 51.3 160 80.6 1.6 48.7 40 19.4 1.5. 3. 15. 14 WH/9 70.2 130 94.1 1.3 29.8 20* 5.9 1.0. 3, 20, 13 1.8. 3, 20. 13

Table II: HGMS TEST RESULTS WITH CLASSIFIED FINES(MURP)

Expt 1.0 Tesla 1.5 Tesla Feed Grade No,. HAG (J3O8(ppa) xwt U3°8 XD xwt Vfl XD (PP«) (ppn) 1 59.33 162 88.6 65.16 154 92.6 108 2 53.88 210 88.7 59.65 199 93.0 128 3 34.36 239 75.2 46.63 207 89.3 109 4 43.32 204 87.2 49.49 180 93.5 95 - 329 -

Table III: RESULTS WITH SURP SAMPLES

Code MAG+MID NMAG Operating Parameters Xvrt (T.rpm.Xsd.lpm) U3°8 %R ER %wt U3°8 %R PPm ppm WH/2 56.4 208 87.3 1.5 43.6 40 13.0 1.8. 3. 20. 12 WH/3 53.8 209 85.3 1.6 46.2 42 14.7 1.8. 2, 20'. 12 WH/4 49.6 194 82.7 1.7 50.4 40 17.3 1.8, 2. 20, 14 WH/5 45.8 206 81.3 1.8 54.2 40 18.7 1.8. 2. 20, 14 WH/6 51.4 234 85.8 1.7 48.6 41 14.2 1.8. 2. 15. 14 WH/10 65.0 185 93.0 1.4 35.0 26* 7.0 1.0, 3, 20. 13 1.8, 3, 20. 13

Table IV-RESULTS WITH RURP SAMPLES

Code MAG+MID t4MAG • Operating Parameters %wt Xwt (T,rpm.Xsd,lpm) U3°8 \R ER U3°8 %R ppm ppm WH/6 38.2 140 65.5 1.7 61 8 45 34.5 1.5, 2. 20, 15 WH/7 41.3 134 65.9 1.6 58.8 48 34.1 1.3, 4. 20. 15 WH/8 36.6 142 65.1 1.8 63.4 44 34.9 1.3. 2, 20, 15 WH/9 46.8 164 75.4 1.6 53.2 47 24.6 1.8, 3. 15. 15 WH/10 54.7 130 77.7 1.4 45.3 45* 22.3 1.0. 3, 20, 13 1.8. 3. 20. 13

* Final reject (NMAG) of two stage operation in WHIMS Table V: LIBERATION CHARACHTERISTICS

MURP SURP RURP xwt 55.8 63.6 71.5 17 18 12 Quartsatlc U~0 (ppm) o 8.3 8.8 9.7 X Disxwt 38.4 31.6 24.1 125 107 83 Micaceous UoOg(ppm) 42.8 26.7 22.1 X Dist Liberated U XWt 5.8 4.8 4.4 6V with UgOgCppm) 940 1707 1395 Magnetite etc. X Dist 48.9 64.5 68.2 xwt 100 100 100 112 127^ 90 FEED U308rppm) 100 100 100 X Dist

Table VI: Feed Size Analysis & Uranium Distribution MURP SURP RURP Particle %D Slze(Mm) xwt XWt XD XWt XD • 208 11.6 8.8 17.4 13.2 5.4 2.5 +147 23.0 13.0 12.5 7.0 12.5 6.4 • 105 21.0 18.0 17.9 18.2 19.0 9.8 +74 15.0 13.8 18.0 7.5 17.7 11.7 + 53 10.6 10.0 13.6 13.6 17.1 12.7 + 37 5.6 7.5 6.3 8.5 10.6 12.9 -37 13.2 28.9 14.3 32.0 17.7 44.0 - 331 -

Fig 1. OISTHI&UTION SC KECOVERV IN DIFFERENT SIZES

2* - m 22 - 1C - • • - ^ X ^ 4 - 2 - i •»-83 -1-7i4 i-J7 I •H47 1-208

Fig. 2. DISTRIBUTION & RECOVERY IN DIFFERENT SIZES /

-3T +3T -»03 t-74 tlOl -H47 •••20*

Fig. 3. DISTRIBUTION & RECOVERY IN DIFFERENT SIZES tim OME : msm/mt/»

+ 7* +IOI t>l47 +2Q« __CA

N.P.H.Padm&nabhan, U.Sridhar and N.K.Rao

Ore Dressing Section, Bhabha Atomic Research Centre. Hyderabad.

Preliminary beneficiation studies were carried out on a small bore-hole uranium ore sample from Tummalapalle. Cuddappah district (Andhra Pradesh). Most of the uranium values occur in this reasonably vast deposit in the form of fine grained pitchblende. The host rock ic essentially dolomitic/ phosphatic lime-stone with small amounts of quartz and shale. The presence of such high amounts as 60-65% by weight of acid-consuming carbonate minerals forbids the adoption of the conventional acid-leaching process for uranium extraction. However, if the acid consuming material in the ore is either removed, or at the best reduced by physical seperation method, with out any significant loss of uranium values, the acid leaching process might still be viable both technically and economically. With this aim,preliminary studies were conducted to separate essentially the carbonates by physical separation techniques.

The ore sample contained 60-65% by weight of carbonate minerals, 10% of apatite and quartz each and about 5 % of pyrite. Radiometric estimations gave the uranium assay as 0.05% U 0 *q. The ore sample was calcined at about 900°C 3 8 for two hours and the calcine was quenched in cold water. The slaked lime formed, is then removed by any one of the methods such as desliming, flotation and dissolution. About 20% of the weight was lost during calcination by the expulsion of carbon dioxide.In the desliming method additional 35% of the weight could be discarded with only about 30% of loss of uranium values. In the calcination and flotation experiment, weight loss was only 20% since slaked lime did not float well. Attempts were made to - 333 -

dissolve the slaked lime, and thi3 way about 70% of the weight could be discarded, with about 20-30% uranium loss. Straight flotation of carbonate minerals with sodium oleate also, gave encouraging results. Other alternate methods like selective dispersion of calcite or selective flocculation of apatite, quartz and pyrite are 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 WEIGHT * °3°8 U3O8DISTN. MESH % ASSAY % + 35 26.9 0.048 30.2 - 35 + 50 17.3 0.045 18.3 - 50 + 70 9.1 0.046 9.7 - 70 + 100 11.7 0.043 11.7 - 100 + 150 5.6 0.041 5.4 - 150 + 200 4.1 0.041 3.9 - 200 + 270 3.0 0.035 2.5 - 270 22.3 0.035 18.3

FEED 100.0 0.043 100.0

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 ores also. (2) 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

Expt.1 Expt.2 FRACTION Assay Dlstn. Assay Dlstn. WT X XU3O8 U3O8 % WT X VJ3O8 U308 X

Sands 41.3 0.1 70.2 35.8 0.12 73.2 Slimes 35.0 0.05 29.8 39.3 0.04 26.8 Wt.Loss 23.7 24.9

Feed 100.0 0.059 100.0 100.0 0.059 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

Expt.3 FRACTION Assay Dlstn. WT % XU3O8 U3O8 % Float 19.6 0.047 16.3 Tails 59.1 0.08 83.7 Wt.Loss 21.3

Feed 100.0 0.056 100.0 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

Expt.4 FRACTION WT % Assay Distn. XU3O8 U3O8 % Solids 36.5 0.12 80.9 Fines 20.7 0.05 19.1 Wt.Loss 42.8 Feed 100.0 0.054 100.0 Solids 57.2 0.095 100.0 Fines

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

Expt.5 FRACTION WT X Assay Distn. X03O8 U3O8 % Float 61 .8 0.03 30 6 Tails 38 .2 0.12 69 4

Feed 100.0 0.060 100 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 % 100.0 Ground Ore Feed U3O8 AssayX 0.115 U3O8 Dlst.% 100.0 I DESLIMING CYCLONE

81.7 Under flow Overflow 0.119 84.6 18<.L3 0.09 Silicate DIRECT LOTATION 15.4 Tails PH = 10 Float 32.2 0.06 Calcite 17.6 REVERSE FLOTATION PH = 5.5 Float Sink 1 j 13.5 0.029 36.0 0.204 3.4 63.6 Apatite Concentrate

Figure 1. Flow-Sheet for Processing Itataia Phospho-Uraniferous Ore 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 with respect to ore dilution? J.L. BHASIN : Supervision of blasting operation is very important. Dilution of ore to the extent of about 10 per cent is 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 a given quantity of ore Increases reducing the production capacity of the mine and hence the mining costs increase with depth*

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 in our plant trials beyond additional 8-10 per cent using Mosabonl tailings.

N.K. RAO t Are you using the aame type of table in all your plants or are you using different types? U.K. TIWARI i More or less same type of tables are used. Of course some modification in the tables have been made from time to time 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 uranium mineral 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 course BMS having very high capcity could be the choice for the first stage.

M.C. BHURAT > Why not apply direct leaching technique for copper tailing?

K.3. KOFPIKER t I would like make comment. In the next session a separate paper is being presented on direct leaching of uranium from copper tailing. Hence this aspect can be discussed at that time*

S. SEN t TV» yOU 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 any control on particle else beoause grinding Is done by H.C.L. to salt their copper recovery circuits.

K.K. DWIVSDY t I feel that either magnetic separation or direot leaching should be followed. By magnetic separation about of 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* SESSION III B

ANALYTICAL TECHNIQUE'S IN URANITTC TECHNOLOGY - II

Chairman : Shri L.M. MAHAJAN (Retd.) i? A R C Reporteurt Shri N. S^AM SUNDAR N?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: -Cx y - 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- Not ITl) MULTIPLIER CRYSTAL

PREAMPLIFIER A.O.C. CONTROLLED 4KUCA

INPUT/OUTPUT DEVICE T OUTPUT

FlG.1. BLOCK DIAGRAM OF Y SPECTROMETRY SYSTEM FCR NP ANALYSIS OF URANIUM

J-Wt-70 Rtv.

FIG. 2. GAMMA SPECTRUM OF URANYL NITRATE SOLUTION •oo

8

»to MM CO •• 70 7t 3 CONCENTRATION OF U IN «••/Litre

790

ISO 170 190 Fit.4 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 is one of the most sensitive instrumental method* of esti- 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 is pipe- tted into platinum dishes specialty made for f luorlaetric work and the solvent is evaporated under an Infra-red lamp. The residue is 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 is cooled and the fluorescence of the resultant bead is 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 filter 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 felt 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 it 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 extraction for fluorimetric determination with the same accuracy as in the case of ethyl acetate aluminum nitrate.

INTRODUCTION

The fluorescence of uranyl compound on irradiation with Ultraviolet light is wall known effect. It was discovered by Becquarel and Stokes in the middle of previous century, since that time the phenomenon has been carefully investigated by many authors. One of the results of their efforts was the discovery of the natural radioactivity and another con- sequence vas the development of a very sensitive method for the qualita- tive and quantitative determination of uranium (Uranium Fluoriaetry) in water, minerals, biological materials etc* Fluotfl metric method for de- termination of uranium in solution is in use right froo the day* of Manhattan Project in U.S.A. The method followed at present is mostly in the line with the one proposed by Grimaldi and further improved by Centanni, 2 Bos* and Oesesa . The procedure followed at Jaduguda for process stream samples involves separation of uranium from sample solutions using ethyl acetate a* an extract ant in presence of saturated solution of aluniniuB nitrate, drying of * measured aliquot of extract, fusing the dried all. quot with sodium fluoride - sodium carbonate flux and than determining the uranium by flnorimetric method. Sine* most of the plant samples are in sulphuric acid medium, a solvent suitable for sulphate medium was fait to be more useful in place of etbylacetate - Aluminium nitrate system. Therefor a it was decided to us* a solvent which can extract uranium In sulphate medium.

Long chain aliphatic amines are ooing used extensively in uranium industry for uranium extraction from sulphurlo acid loach liquors. One of the popular amines, Alav.Jie-336 was used in our B&D studies for uranium at- traction from Jaduguda leach liquor. Since it was found to be a good ex- tract ant for uranium in sulphurlo acid medium, the saae solvent vac , solsctod for extraction of uranium in process stream senples for flnori- aotric estimations of uranium and the estimations were carried out suoo- - 358 - easfully. Studies were farther extended for assaying niU feed samples sod sample• froa other sines of Singfrhhai in place of T.B.P, extraction speotrophotoaetrio aethod. The results obtained are dealt with in this paptr.

(a) Seageats and Ch—deals

(i) Xtamlno-336 (l* VA solution) t~ ID ml of alaalno-336 was diluted to 1 litre using IR grade bensene. The diluted solvent was washed twice with 100 al portion of water of pH 1.0 and was filtered with whataaiw40 filter paper and stored in a bottle.

(11) Flux Mixture!- 4il Mixture of sodlua oarbonste aid sodlua fluoride .

(ill) Standard ttraniusi Solatloni- (lO^ng/al) in sulphuric acid •edium.

(b) aaaple Preparation*. 1 ga of the powdered ore/Tailings saaple was digested with 50 al of an aold aixturc oontaining 5 al of nitric add, 5 al of sulphuric aojd and rest water for 1-4 hours on a hot plate and evaporated to dense sulphurie acid fusing. The fusing was continued for SO to 45 ainutee and then the beaker was ooolad and SO al of water was added earefully end was boiled for 5 to 10 alnutes. ifter cooling It was filtered and the voluas was aede up to 250 al la cav> of ores and 100 al in cass of tailings*

(o) Uraniua &ctractlon froa Solutions*-. Ursniua was extracted with Alamlne-356, as an extract ant in benaene diluent,' The effect of

other diluents such as, cgrolohexane, ether and etbrl aoetst«f con- oentration of extraetant, pH of the aqueous phase, tiae end phase ratio on the extraction of uraalua were studied* The used solvent - 359 - was regenerated by washing twice with 2.5 percent sodium carbonate 8olution(l litre of used solvent with 100 ml of 2.5£ sodiun carbonate solution) followed by one wash with 1.0 pH water and finally with dis- tilled water.

(d) Procedure for analysis!- A suitable aliquot of the solution samples of the plant and of the ores and tailings as prepared above was taken in stoppered extraction tube and volume was made upto 10 ml keeping the pH in 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 settle for complete phase separa- tion, 0.1 ml or 0.2 ml of the extract was pipetted out into the f luorimetrio platinum dishes and the solvent was evaporated under IB. lamp. 0.4 to 0.5 ga of sodium carbonate sodium fluoride flux mixture was put into the dishes and was fused at 800°C for 3 minutes. After cooling the intensity 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 Intensity of fluorescence measured were oompared to know the unknown concentration.

The effect of addition of interferences on the accuracy of uranlua es- timation by using Alamine-336 ware also studied. Experiments were also conducted for supresslng the interferences.

RJBULTS AMD DISCOSSIOMS

1. Test for the Suitability of Jbctractant Concentrations Uranium was extracted from the solution containing different amounts

of l^0Q using & alamine-336 in benaene. A. plot of f luoriaetrlo readings against quantity of UgOg in solution is shown in ?ig.(l).

The straight line in Fig.(l) shows that upto SO ug of 1%0Q oan be extracted 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 one percent alamine-336 in benzene and by using ethyl acetate abminiua nitrate were analysed fluorimetrically and the results are compared in Tabla-I.

TiBLE - I

Samples (ga/1) {After extraction with j ifter extraction with Il£ alamine-336 in benzene I ethyl acetate AL- I f nitrate medium

Ion exchange Barren 0.0052 0.0054 Plant total Barren 0.0033 0.0032 Barren Diversion 0.010 0.010 Secondary filtrate 0.0137 0.0130 Break through sample 0.0013 0.00135

The results show that there is a very good agreement in tho results obtained which confirms the applicability of alamine_336 as an extract ant in place of ethyl acetate aluminium nitrate system to extract uranium from 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 were analysed several tl»e by using ilsmlne-336 in bensene and by ethyl acetate w"1'—<«fc»» nitrate for the extraction of uranlun. The results are 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 No. of Value obtained I Average value] Standard deter, g/1 IU>8 deviation mination

A 7 0.00196 0.00176 0.00184 O •000071 B 7 0.0020 0.00176 0.00186 O•000085 Barren . diversion A 5 0.0310 0.0290 0.0299 0.00081 B 5 0.03ID 9.0285 0.0293 0.0010 Secondary - pulp A 7 0.034 0.031 0.0323 0.00106 filtrate B 7 0.034 0.031 0.0320 0.00106

33SSXSSS = s s s s a

From the result* it is clear that the precision of analysis in case of alanine-336 extraction is batter than that of ethyl acetate aluainiua nitrate systea.

(4) application of ^ine Extraction in the Analysis of Tailings of the giant.

The tailings samples were analysed for uraniua f luoriaetrically using alenine-3'36 in bensene and ethyl aoetate-alaainlw nitrate as an extractant. The results obtained by two different routes oonpared in Table-Ill. - 36? -

- ni ANALYSIS Oy TAILINGS SAMPLES USING TMD DIFFgtEHT EXTRACXANTS

Samples *°3°8 as an I Using AL-nitrat© and Ethyl- Sxtractant I Acetate as an Sxtractant

0.0048 0.0044 0.0042 0.0042 Leaching Pachuca 0.0040 0.0045 Tailings 0.0040 0.0042 0.0041 0.0043 0.0042 0.0041

0.0075 0.0080 0.0077 0.0078 Tailings pachuca 0.0078 0.0081 (Final Tailings) 0.0078 0.0076 0.0077 0.0076 0.0084 0.0079

S S 3 B From the results In the table it can be seen that the ralnes obtained by 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 plant and uranium ores from Narvapahar and Turandih were analyael for uraniua f luorlaMtrically using alamlne-336 in. bensene as an extractant and spectrophotoaetricalljr using T.B.P. extraction method. The results of OOP analysis by two different methods are oompared In Table-IV and of Harvapahar and Turamdih ores are compered in Table-V. - 363 -

TJPLB - IV OOMPAHISION OF COP ANALYSIS BY TWO DIFPBtHfl METHODS

Date *U3°8 I Date *D3°8 By Colorine-j By fluorl- By Colori- |By f luorime- tric method | metric method metric [tric method L method 1 0.060 0.061 16 0.061 0.057 2 0.059 0.058 17 0.052 0.C52 3 0.058 0.060 IB 0.057 0.055 4 - - IS 0.060 0.060 5 - - 20 0.058 0.C58 6 0.062 0.062 21 0.052 0.062 7 0.C59 0.058 22 0.055 0.057 8 0.060 0.C60 23 0.C55 0.055 9 0.057 0.C56 24 0.052 0.053 10 0.058 0.057 25 0.057 0.066 11 0.060 0.061 26 0.057 0.057 12 0.060 0.058 27 0.063 0.059 13 0.085 0.083 28 - mm 14 0.061 0.059 29 0.054 0.057 15 0.057 0.057 30 0.052 0.062 31 0.048 0.048

s s a s s - 364 -

TiBLS-V

ANiLTSIS OF A FBf ORB SiMPLBS FflQM SINGHBHUM iRBA BI TWO PIFFSajgHT MPHODS.

S aa p1 e *°3°8 By Colorimetric proposed Fluorimetric TBP Attraction I mtmethof d

Narva - A 0.0470 0.0475 Turaadih - A 0.059 0.056 • - B 0.050 0.050 » - C 0.045 0.045 • -D 0.042 0.042

It ia clear from the above result a that the values obtained by the two different aethode are aore or laoa same confirming that fliiorlaetric method for uranim eatiaation using alajdne-336 In benaene la a goou substitute of spectrophotoaetric method ualng IBP extraction for the uranivai 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 In the fluorimetric eetlaation of uranim ualng 41amlna-536 in bensene as an extractant ware atodied and the results are plotted la Fig.(2) and Fig.(5). Frm Fig. (2) it can be aeon that Ca lnterfers In uranlwi estimation in amlne ayatem of extraction but upto 1.0 g/1 of Ca can +2 be 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) it can be seen that Chloride does not interfere upto 2.5 g/1. If the concentration goes above it interferes very heavily. If Chloride is present, it should either be brought dovn to 2.50 g/1 level in 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 fluorlaetrio method is most suitable for plant leach liquors as well as for solid ores and tailings samples. The preoislon and accuracy of the method are quite satisfactory* By using this method there will be saving of chemicals Ilk* Aluninlum nitrate, Anraonlum Nitrate and Ferric nitrate. The estimation time of solid samples involving TBP extraction will b* reduced, Tha consumption of solvent will be very snail because the same solvent aft*r washing can b* reused for several cycles*

The authors wish to thank Chairman ft Mur-m^ng Director, Uranium Corporation of India limited for his keen interest in the Besearch activities of CRfcD Department,

1. Grlnaldi F.S. t- " Collect«d paper on method of Analysis for Uranium and Thorium ". Ooo Survey Bull 1006 (1954) U.S. Govt. Printing Offioe, Vaahington.

2. Centanni P.A., Boss A.M. and Oesesa M.A. " Pluorimetrio Determination of Uranium «. Anal. Chen. 28, 1651 - 1657 (1956). - 366 -

U3O8 EXTRACTION USfNG ONE PERCENT ALAMINE 336 IN BENZENE

4000

o 5 3000

Ml at o 2000-

JO 20 J0_ .40 5Q

Ofi IN SOLUTfON,^

" Fl<3. 1 - 367 -

INTERFERENCE OF Th, Ce AND NO3 IN AMINE 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*0 CONCENTRATION, no. z INTERFERENCE OF Cf IN AMINE EXTRACTION OF URANIUM

|100 O O 0 oK 90 c 80J x 70

§50! £ 40 30 2Q

O. f 00 ro ido CHLORIDE ION CONCENTRATION

FIG. - 369 -

DETERMINATION CP URANIUM IN SEA WATER BT ABSORPTIVE DIFFERENTIAL PULSE VOLTAMETRT

R.N. Kbandekar and Badha Ragbunath Pollution Monitoring Section Bhabha Atomic Research Centre Trombay, Bombay 400 065.IIMA

An adeorptiTe a tripping voltammetric procedure for direct determination of trace quantities of Uranium in eeawater has been described. Optima 1 conditions include pfl 6.7, 2 x 10* M 8-hydroxy quinolint (Ozine) and collection potential of -0.4V (Vs AgAgcl) at banging aercury drop electrode. With controlled adsorptive accumulation for one Bin. a detection limit of 2.8 z 10 M Uranium is obtained. The response is linear up to 7 i 10 M Uranium and the relative standard deviation at 4 z 10 MU is 11.5J*. The effect of possible interference from other metals has been investigated.

IHTRODUCTIOM

Detexmlnation of uranium in sea waters is of interest because the element is used for the production of eneny in nuclear reactors. The contribution fvom nuclear facility, being small, is often masked by the relatively high variability of uranium concentrations in coaetal sea water. The stable oxidation state is uranium (vi)in oxygenated waters and is mostly present as uranyl ion which is 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 it 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 of uranium in sea water. It would obviously be useful to be able to determine the concentration of uraoiua directly in the sea water without prior cherical separation. Recently oathodic stripping voltammetry technique was developed for the determination of uraniun in natural waters* ' . This paper presents a sensitive and rapid adsorptive stripping procedure for the determination of uranium in sea water using 8-hydroxy quinoline (Oxine).

Apparatus x PAR 174-4 polarographio analyser with PAR 303 banging mercury drop electrode (HMDS) and PAR 305 electro magnetic stirrer were 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) was prepared and diluted as per the requirement. An aqueous stock solution of 0.1 H oxine was prepared in 0.25M HCl (Analar, BOB) and diluted with distilled water. A pH stock solution of IM PIPS (piperaiine-I-H'bis 2-etbane sulphonlc acid) mono sodium salt *nd 0.5M VaOH (Arlstar, BIB) is also prepared. An aqueous solution of 0.1M EDTA was prepared from its sodium salt and was adjusted to pH~7 by laOH. Sea water samples were colleoted from coastal places in India. On collection in precleaned polythene bottles, they were acidified with HCl so that pB of water is nearly 2. Before measurement sea water samples were filtered through 0.45 .urn membrane filters.

Procedure t A 10ml of sample aliquot was taken into the voltammetric cell and 0.2ml of 0.001M oxine solution and 0.1ml of PIPES pH buffer were added. After deaeration of the solution for 8 min. by purging highly purified nitrogen, a fresh mercury drop was extruded. The Sjtlrrer was started and electrolysis was done for - 371 -

1 min. at —0.4V. The solution was allowed to become quiescent and the cathodic scan was carried out in differential pulse mode with scan rate of 5 mVs~ and the sensitivity of 500 nA(full scale). The measurements were repeated after 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 electrolyte, adsorption potential and adsorption time, scan rate and biological buffers vis HEPES and PIPES.

Biological Bufferes j Effect of biological buffers on the peak current of uranium was studied. The voltamnetrio cell 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 sin. It 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 is 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 at oxine concentration of 2.5 x 10~ M. Por analytical purpose, therefore, the axlne concentration of : 2 x 10 M was used. It was observed that pH between 6.5 and 7 is quits adequate tor the above oxine concentration. The peak current increases with the increase in pH, however it 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 —1 remained almost constant for seen rates, 1 to 5 mYs and then decreased with the increase in scan rate. For all 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 is -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.

It 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 is linear 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 sensitivity dropped considerably. The sensitivity 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 respectively. These peak potentials are well separated from that of uranium except for Fb and Cd. However it was observed that they do not interfere with estimation of uranium due to their lower sensitivities and the concentrations in sea water. Both Pb -4 and Cd can be masked completely by addition of 10 M EOTA to the sample where as Uranium peak is not affected

Several sea water samples collected from different location mostly around Bombay and few other places in India during Jan.1966 to Uarch 1989 were analysed for uranium using the above standardized procedure. The uranium concentration varies from 0.95 to 3.95 juj l"' with the average value of 2.36 • 0.97 )igl~1 (Table 1) the average value is ID close agreenent with the values obtained by other workers ' .

REFERENCES

1. H. Ogata, N. Inoue and H. Kalibana, (1971), Nippon Genshirycku Gakkaishi, 13, 560-564

2. K. Saito and T.Miy&uchi, (1982). J. Nucl. Sci. 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. Sci., 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. Pillai 1 Beport BARC/I-973, DAE India (1979). - 375 -

TABLE I

URANIUM IN SEA WATER

Cone.of Uxanium S.No. Place of Collection ug l" water 1. Kanyakumari, Tamil Nadu 3.57 2. Kalpakao, Tamil Nadu 1.90 3. Haei All, Bombay 3.81 4. Apollo Pier, Bombay 3.95 5. Cirus, Bonbay 2.14 6. Thane Creak, Bombay 0i95 7. Tarapur Atonic Power Station 1.07 8. Cirus, Bombay 1.67 9. Gateway of India (I) Bombay 2.66 10. Gateway of India (II), Bombay 1.90 11. Band Stand, Bandra (I) 1.07 12. Band Stand, Bandia (II), Bombcy 2.36 13. Bandz* Bombay 2.38 14. Versora (I) Bombay 2.62 15. Veraora (II), Bombay 3.09 - 376 -

DIFFICULTIES IN PREPARING A STANDARD SAMPLE OF URANIUM METAL HAVING TRACES OF NITROGEN

R.S.D.TOTEJA, B.L.JANGIDA, M. SUNDARESAN Analytical 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 N Number 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

A- CONE t SOCKET 8- SILICA REACTOR TUBE C - URANIUM PELLET D - SMALL FURNACE E - OIL MANOMETR 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 is estimated in the tailings effluent after neutralisation with line. Since after neutralisation at 10.00 pH, very little Manganese is left in the tailings effluent, a very efficient method of estimation is required. Foraaldojdme^ method is currently being folloved for the estimation of Manganese spectrophotonetrically in the highly basic solution, At high pH (above pH 10) Manganese content in tailings effluent is so low that it is beyond detection Halt of spectro- photcnetric estimation, ilso, at low pH Formaldoxima method gives high results due to the lnterferenoes of Copper, Iron, Draniisi and VanadiuD. Therefore to analyse Manganese, 1CP-JLES has been used which is very sensitive instrument for low concentrations. It was possible to estimate Manganese upto ppb lerel with 96 - 102* accuracy In plant tailings effluent samples In IC*-AB.

IHTRODUCTIOM

In Uranium Ore prooesalng plant, in order to maximise Uranlun extraction, a suitable oxldant must be added in the ore slurry during add or alkaline leaching. In add leaching process generally NaClO,, 0 or MnOg is used as an oxidant. In Jaduguda Uraniw plant, which processes more than 1000 tonnes of ore p»r day, pyrolnsite is being used as an oxidant, *ich oxldlsts tetraralent Oranlw to hexa-valant state through Ferrous - Ferrio cyle. The barren solution containing a considerable quantity or Manganese to the tune of 1.0 gm/1, after neutralisation to pH io to precipitate msnganess and other elements, is sent to the tailings pond. The discharge of the tailings pond generally goes to - 383 -

the river. During neutralisation manganese is precipitated as Mn(OU)^

or finely divided MnOg. Mn(0H)2 or Mn(0H)3, if 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 is the aquatic pollution since the toxic elanent is 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 al Cone HND5 and volune was made up to 1000 al by adding distilled water. The solutions of required concentrations were made up by diluting the stock solution,

(C) Plant Sappiest The staples of plant tailings after neutralisation with lla* 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/min 4rgon flow I Sample gas - 1 I/mln | iiDdlllary _ 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 to 25 ppa. Maasurejient of tailing* 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 tailing* 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 in the calibration curve of Manganese as shown in Fig.I. It is seen from the calibration curve that manganese can be estimated from 0.1 ppn to 25.00 ppn. Concentration of Manganese in the tailings effluent were estimated both spectrophotometrically and in ICP-JLSS. Results are compared in Table-I. It is observed that both in the ppb and ppo level, manganese can be estimated with 95/t to 102* accuracy in ICP-AES.

TiBLB- I

SI. 1 ifave length! pH of Cone of * Cone of * No.t an Tailings Manganese Manganese I Accuracy Effluent Spectrophoto- in IGP-ABS I netrically ppa I

1 257.61 9.5 0.120 0.121- 100.8 2 257.61 9.7 0.158 0.161 101.9 3 257.61 9.7 0.130 0.128 98.5 4 257.61 9.5 0.360 0.360 100.0 5 257.61 6.5 212.0 208.4 98.5 6 257.61 11.1 0.0076 7 257.61 10.5 0.0056 8 257.61 3.6 630.0 620.0 98.4 9 257.61 4.2 416.0 410.0 98.6 10 257.61 9.9 0.067

3 s : 3 3 3 3

Interferences!

The effect of interferences which are predominant in the Jfar&aldoxlae method were studied in ICP-AES with known concentration of Manganese, The results are given in Table-II. It is seen from T«ble-II that Calcium which is most predominant species in the neutralised «ffluent can be toleratod upto 350 ppa and Iron, Uranium, Vanadium and Copper upto 40 ppa. - 386 -

. n EFFHCr OF IMEStFHUBfCBS

Vaye length} Standard J Interfering Cone of Inter- Measured cone m >nc3ntrstion| element | fering element B 9 of Ma i* of j PfB uaganese ( i I I" Ppl .1 i I 257.61 1.00 10 1.01 • R • 20 0.99 R H R 30 0.38 R R R 40 1.02 R • H 50 1.04

257.61 1.00 0 10 1.01 • fj • 20 1.00 • R • 30 1.01 R • R 40 1.01 • • R 50 1.06

257.61 1.00 V 10 0.98 • R R 20 0.99 • R • 30 0.99 R R R 40 1*002 R R R 50 1.004 257.61 1.00 Ou 10 1.01 R R H 20 1.00 R • R 30 1.02 R R R 40 0.98 R R R 50 1.06 257.61 1.00 Ca 50 1.002 R • R 100 0.99 R R R 150 0.965 257.61 1.00 * 200 1,00 • N R 250 0.99 ft R R 300 0.99 II R H 350 m 0.96 w * R 400 0.97 • R 450 0.96 • S3 3 a * a a - 387 -

CONCLUSION

Manganese vas estimated down to the ppb level in tbe neutralised tailings of Jaduguda Oranium plant in the ICP-AES at the wave length of 257.61 m. The results were very much precise and accuracy varied from 95 to 102*.

ACKNOULEDGBiaC

The authors wish to thank Chairman & Managine Director, Uranium Corporation of India Limited for his keen interest in the research activities of C.R.&D Department.

RgRBMCBS

1. Colorimetric Determination of Traces of Metal - E.B. Sandel. Inter Science Publications INC, New lork.

2. Rankana K and SahaaaTH.G. 'Geochemistry', Dniv. of Chicago Press, Chicago, Illinois, USA (1950) P.640.

3. Studies on Specie* variation of Manganese In Uranium Processing and natural environment - M.Sc Thesis (Chemistry) Shree S. Venkataranan, Boabay diversity 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. Baylls, J.R. Jour. A.U.U.A 32 (1940) 1753 quoted In C.A. 36 (1941) 545.

6. awards G.P. Jour. M.E.W.W.A.6V1947.260

7. Neumann, R. , Z. Gesund neltstech. St^tehyg. 25/1933 163, quoted in 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 detail under different experimental conditions. However not sufficient data is available for the voltammetric uranyl studies at different metal wire and metal amalgam wire electrodes. Uranyl reduction, therefore, was tried at gold wire, as well as, gold amalgam wire indicator electrodes, employing the three electrode system where platinum wire electrode is auxiliary and molybdenum wire is reference electrode. The study was carried out in acidic, neutral and alkaline media, and in buffered and unbuffered solutions as well. In acidic medium at gold indicator electrode only a single curve was obtained in different acids and buffers. The highest current of 83 jiamp is obtained in 0.1 M hydrochloric acid for 1.0 mM uranyl concentration. No disproportion of U(W couli be detected at gold indicator electrode. However in contrast to acidic media, two peaks were obtained in 0.1 M each of potassium nitrate and potassium chloride medium, current concentration linearity being obtained for both peaks. The highest currents were obtained in 0.01 M potassium chloride, being 75 and 335 jiamp respectively for 1.0 mM uranium concentration.

Similar results were obtained at S^ld amalgam electrode though the current heights'obtained in acidic media at the amalgam electrode are considerably smaller than those at gold electrode. In neutral media, the results obtained at gold electrode are comparable to the results obtained at amalgam electrode. Thus amalgam electrode does not behave as mercury electrode. Comparision of the results obtained for these three electrodes is discussed. - 390 -

VOLTAMMETRIC STUDIES QF U(VI) REDUCTION INTRODUCTION

The polarographic study of uranyl ion is extensively 1 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 reduction 2 was studied at hanging mercury drop electrode, carbon or glassy 3.4 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 wire 6 as a reference electrode . Gold wir* electrode was cleaned by - 391 - cathodisation in 1 : 4 sulphuric acid for ten minutes. Molybdenum 6 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 2nd Interscience Publishers, New York,1952.pp 462 et seq 2. J.Ferreria, S.Batstachaves, M.Fatima, A.Abrao Chem. Abstr. Vol 108 (1988) 153849 3. K.Izutsu, T.Nakamura, T.Ando Anal. Chim. Acta, 152 (1983) 285-8 4. K.H.Lubert, M.Schnurrbusch ibid, 186 (1986) 57-69 5. C.A.Harte, B.P.Sanchez Chem. 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/Mo Voltage Scan Rate : 40 mV/sec

1 2+ Sr.No. !UO Cone.1 Supporting Electrolyte I Ep 1 iP I 3 1 I Remarks 1 ! mM t M t V 1 jiamp 1 1 0.5 1 Acetic acid, 0.1 ! -0 34 1 20.5 1 Peak !

2 ! 1.0 i * H 1 -0.37 36.0 1 " I 3 ! 0.5 1 Acetic acid, 0.1, pH 3 1 -0.29 24.5 Peak !

4 1.0 H 1 -0.32 47.5 it 1 5 0.5 Acetic acid, 0.1, pH 4 -0. 32 25.5 Peak !

6 1.0 M n 11 -0.31 47.0 I _—__— _ 7 0.5 Acetic acid. 0.1. pH 6 -0. 31.5 Peak ! 36 ______— -— 8. 1.0 n H -0. >< 1 26 67.0 1 —— — 9 0.5 Disod. EDTA. 0.01 pH 6 -0. 12.0 S-type 1 22 10 1.0 tt II -0. 24 26.0 ! 11 ! 0.5 1 Nitric acid, 0.1 -0.15 15.0 S-type I

12 I 1.0 I n H -0. 12 30.0 II l 13 ! 0.5 ! Hydrochloric acid, 0.1 I -0.18 ! 39.0 Peak !

14 i 1.0 ! *• H 1 -0. 13 1 83.0 ! M 1 , - 398 -

TABLE II URANYL REDUCTION J NEUTRAL AU£ ALKALINE MEDIA AJ£ GOLD ELECTRODE Apparatus Electroscan-30 Electrode System : Au/Pt/Mo "oltage Scan Rate 40 mV/sec

I 2+ 1 Sr.No. !UO Cone. 1 Supporting Electrolyte Ep ip ! Remarks ! 2 t mM 1 M V jiamp 1 1 0.5 1 Acetic acid. 0.1, pH 7 1 -0.39 26.5 Peak

2 I 1 .0 H H It 1 -0.38 34.0 3 0 .5 Pot. Nitrate, 1.0 -0.20, -0.71 34 .0,170.0 Both Peaks

4 1.0 M ft -0. 22,-0.62 80 .0,265.0

5 0 5 Pot. Nitrate, 0.1 -0. 84 176.0 Peak 6 1 0 -0.85 290.0 n 7 0 5 Pot. Nitrate, 0.01 -0. 20,-0. 85 28 .0,156.0 Both peaks 8 1 0 II -0.23,-0.90 55 .0,320.0 ii 9 0 5 Pot. Chloride 0.1 -0. 20,-0. 80 40 .0,154.0 Both peaks 10 ! 1.0 ! II -0. 30,-0.92 90 .0,315.0 11 ! 0.5 ! Pot. Chloride 0.01 -0. 22, -0.90 46 .0,154.0 Both peaks

12 » 1.0 H II 1 -0. 31,-1.04 ' 75 .0,335.0 I 1 1 — ! For alkaline solution (Disod. EDTA, 0.01 M, pH 8 ) no proper graphs. - 399 -

TABLE III

URANYL REDUCTION ACIDIC MEDIA A£ GOLD AMALGAM ELECTRODE Apparatus : Electroscan-30 Electrode System : Au(Hg)/Pt/Mo Voltage Scan Rate : 40 mV/sec

2+ Sr.No. UO Cone. Supporting Electrolyte Ep ip Remarks 2 mM M V jiamp

1 0.5 Acetic acid . 0.1 -0.38 11.2 Peak 2 1.0 -0.37 19.6 » 3 0.5 Acetic acid .. 0.1. pH 3 -0.90 6.4 Peak

4 1.0 it It ft -0.96 13.0 it 5 0.5 Acetic acid , 0.1. pH 4 -0.40 13.0 Peak

6 1.0 H n II -0.40 25.5 it 7 0.5 Acetic acid , 0.1. pH 6 -0.42 16.0 Peak 8 1.0 •I -0.35 42.5 9 0.5 Nitric acid , 0.1 -0.72 15.5 S-type 10 1.0 II -0.70 31.5 - 400 -

TABLE IV URANYL REDUCTION U£ NEUTRAL AND ALKALINE MEDIA AT. GOLD AMALGAM ELECTRODE

Apparatus : Electroscan-30 Electrode System : Au(Hg)/Pt/Mo Voltage Scan Rate : 40 mV/sec

2+ Sr.No. UO Cone. Supporting Electrolyte Ep ip Remarks 2 1 mM M V jjamp 0.5 Pot. Chloride, 1.0 -1.07 174.0 Peak 2 1.0 -1 .05 335.0 "3 0.5 Pot. Chloride. 0.1 -0.26,-1.08 24.0,182.0 Both peaks

4 1.0 n i» -0.29,-1.12 55.0,400.0 - 5 0.5 Pot. Chloride, 0.G1 -1.10 142.0 Peak 6 1.0 n -1.18 300.0 7 0.5 Pot. Nitrate, 0.1 -1.00 158.0 Peak ^

8 1.0 H ft -1.04 295.0 tt 9 0.5 Pot. Nitrate, 0.01 -1.13 144.0 Peak 10 1.0 -1.19 315.0 Peak 1

For alkaline solution ( Disod. EDTA. 0.01 M, pH 8 ) no proper graphs. - 40'. -

Apparatus; Electro scan - 30 Electrode System ; Au/pt/^4o 7- Voltage Scan Rate : 4Omtf/sf5 ~ SGfBflmp Current S@nsHiv

o I J- z 3

or 3

- 0.6 -0.8 - )'V - i •* - '••« - 0-2 4- 0-2 0.0 VOLTAGE Rg - 1 REDUCTION IN 0,01 M POTASS.UM NITRATE AT GOLD WIRE ELE C T RODE URANYL - 402 -

Session III B DISCUSSIONS

Paper No. 2

S. GANAPATHY IYER s In normal fluorimetrio method the calibration graph is constructed for the rarge 0,01 - 1ug U. In the projected calibration curve the concentration range is given as 10-50ug. I feel there is acme oversight in this*

A.B. CHAKRABORTY t The range shown is what if ^resent in 1Cml of the solvent phase. Since 0.1ml of this extract is taken for actual fli'orimetric analysis, the range for calibration may please be read as 0.1ug to 0.5ug U,Og. Since most ' the samples analysed by fluorimetry in our laboratory fall in itua range of uranium concentration, the calibration graph given is also for' this range. The amlne concentration (1£ antne) has been found to be suitable tor this range*

Paper Wo, 5

B.L. JANGIDA t What la the significance of determining of Mn in uranium tailIng3?

JOYDEB RAY t During neutralisation, manganese Is precipitated

as Mn(0H)4 or finely divided Mn02Mn(0H)2 or Kn(0H)?. If present aa Mn0g.Mn(0H)2 or Mn(OH), it gets converted to Mn(0H)4 by atmospheric oxidation which la not stable in aqueous solution* In course of time Mn may affect the environment through the media of water for domestic purposes, and indirect assimilation through food stuff are the common mode of health hazard which affects mainly the central nervous system. Therefore a statutory level 0.5 ppm in the neutralised tailings plant effluent has been recommended*