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Sedimentary Ores-Ancient and Modern (Revised), edited by C. H. James. Proceedings of the 15th Inter-University Geological Congress. University of Leicester, December, 1967. Special Publication No.1, Department of Geology, University of Leicester, February, 1969. Obtainable from: Leicester University Bookshop, May's Walk, Leicester LE 17 RD. Price: 42s. Postage and packing, 3s. 6d. This volume should be of interest to economic geologists and students in this field. A number of illustrated papers on the genesis of ores in sediments are presented; the authors include experts such as Prof G. C. Amstutz, Dr K. C. Dunham, Prof D. Williams and W. G. GarIick. The papers deal mainly with sulphide deposits, but the late Prof J. H. Taylor's contribution concerns ores of iron and manganese. Brief references to the Witwatersrand ores represent the only discussion of South African deposits. With the notable exception of W. G. Garlick's paper on Zambian copper deposits, the Abstracts given are inadequate, necessitating thorough reading of the papers to extract the important matter.

Colloquium and General Meeting were held at Kelvin House on 19th May, 1971, the theme being 'Modern trends in coal mining practice.' Mr V. C. Robinson (President) was in the Chair. The Colloquium was attended by 205 delegates and was opened by the President at 9 a.m. OBITUARY The President: 'It is my sad duty to announce the death of C. C. Cullen, Fellow, who joined the Institute in 1946 and passed away on 15th March, 1971. As a mark of respect to the memory of the deceased and in sympathy with the bereaved I would ask you all to rise and observe a few moments silence'. ELECTION OF SCRUTINEERS The President: 'In terms of Clause 9.4 of the By-Laws I now call for the nomination of seven Corporate Members to act as scrutineers of the ballot for Council'. Messrs V. C. Robinson, Prof D. D. Howat, Dr J P Hugo, D. G. Maxwell, J. K. E. Douglas, R. J. Adamson and P. Lambooy were nominated by P. W. J. van Rensburg and seconded by P. A. von Wietligh. The President: 'Are there any further nominations? If not I declare these gentlemen elected.' MEMBERSHIP The President: 'I have much pleasure in announcing that the names of candidates, having been published in accordance with By-Law 5.2.2, Council has elected them to membership of the Institute in the following grades: Members: Douglas Edward King, Bryan Rudolph Scott, James Gregor Phimister, Anthony David Walters. Graduates: Cornelius Johann Muller, Marthinus Stephanus Mulder. Associates: John George Everson, Colin Robert Llewellyn Davies. Students: Francois George Enslin Beetge, Ewald Boshoff, Christiaan J. Cloete, Christopher John Davies, Cameron George Elvin, Kenneth Eric Field, Johannes Louis Fourie, Hans Gert Gastrow, Peter John Ledger, James Robert William Lindsay, Kynaston Lloyd McDonald, Richard Peter Mohring, Arthur Leslie Painting, Matthias Arthur Pascall, Ian Neil Sinclair, Donald Allan James Ross-Watt. MEMBERS TRANSFERRED TO A HIGHER GRADE From Member to Fellow: Peter Norman Roberts. From Associate to Graduate: Alfred Eric WaIter Fletcher. From Student to Graduate: Wilfried Pierre de Villiers, Edward Emile Eichenberger, William Alan Nairn, Lodewyk Johannes de Jager, Colin Henry Obray, John Grenig Rees. COLLOQUIUM The President opened the proceedings and introduced Mr N. W. S. Schumann as the overall chairman for the day. The following papers and contributions were presented: FIRST SESSION: Chairman R. C. J. GOODE. Paper: 'Problems encountered in the operation of a new colliery' by P. M. C. Wilson and A. A. Oakes. Contributors: D. J. Moloney, G. R. Canny, I. G. Evans, R. B. MacGillivray, Dr F. G. J. de Jager, M. J. Deats, and N. Zolezzi. R. B. MacGillivray Mr President and gentlemen, I would like to congratulate the authors on presenting a most interesting paper on a very important subject. After several attempts at finding suitable reserves to extend the life of Welgedacht Exploration Company's Utrecht Colliery a block of coal rights was acquired south west of the town of Utrecht. A study of these reserves showed that they would enable a satisfactory return on capital invested to be achieved. It was decided therefore to open up what is now called the Zimbutu section of Welgedacht Exploration Company, to produce 41 000 metric tons of coal per month. The decision to open this section was taken in August, 1968 and due to certain allocation requirements it was necessary to bring it into production by the 1st of January, 1970. Fortunately a fair amount of statistical data had been collected prior to August, 1968 from the old Utrecht Colliery in anticipation of opening a mine and a fairly detailed rechnical report had been prepared. As a result the dead line was only missed by four days. Main features of the Design A feasibility study was carried out on three possible shaft positions and the final position was selected from these three on economic grounds. As the coal seam at the selected shaft position was only 21 metres below the surface it was decided to sink two twin incline shafts to enable rapid establishment of through ventilation and alternative hauling arrangements for quick development. This also allowed the final vertical ventilation shaft to be placed at some distance away to avoid the nuisance of fan noise. Second outlet requirements are also very well met by these twin inclines. Again on the basis of economics and also due to the uncertainty of roof conditions it was decided to use hand loading methods underground. However, to allow for future possible conversion to mechanised loading, should this become economical through rising wages, a conveyor belt was installed in the hauling shaft. This conveyor belt is fed from an endless rope haulage in the east companion by means of tipplers, a bin and a feeder. The main development heading was left clear for the future installation of a conveyor belt which would afford easy conversion to mechanisation.

By C. Chapman, M. G. Reading, M. P. Amsden

Computer control of flotation at the Ecstall concentrator By M. P. AMSDEN, C. CHAPMAN, and M. G. READING Ecstall Mining Limited, a subsidiary of Texas Gulf Sulphur Company, operates the Kidd Creek Mine, a 10000 tons per day base metal producer, near Timmins, Ontario. Two different types of ore are mined and milled: copper-zinc and silver-lead-zinc. At the highly automated Ecstall concentrator the ores are processed by selective flotation in three individual circuits. One of the most important metallurgical tools for process control is a sample system incorporating continuous on-stream X-ray analysis. Recently this system was automated further by the addition of computer control of reagents for two copper-zinc flotation circuits. The computer programs, utilizing optimum seeking methods and smelter contract terms, have increased copper recovery by 1,1 per cent, zinc recovery by 1,4 per cent, and have reduced reagent costs by 16 per cent. Although not expected, copper and zinc concentrate grades were improved slightly as well.

The Annual General Meeting of the Institute was held in Kelvin House, Johannesburg, on Wednesday, 25th August, 1971, at 4.15 p.m. Mr V. C. Robinson (President) was in the Chair. There were also present fifty one Fellows, twenty one Members, two Associates, two Students and fourteen Visitors, making a total of ninety one. The President declared the meeting open at 4.15 p.m. OBITUARIES The President: 'It is my sad duty to announce the death of the following members of the institute: H. J. G. C. Arndt, Life Member, joined 1933, passed away 11th May, 1971. J. H. A. Diering, Fellow, joined 1921, passed away 24th June, 1971. L. O. W. Leimer, Fellow, joined 1960, passed away 25th June, 1971. E. A. Meltzer, Fellow, joined 1941, passed away April, 1971. As a mark of respect to the memory of the deceased and in sympathy with the bereaved, I would ask you to rise and observe a few moments silence.' MINUTES The President: 'May we confirm the minutes of the General Meeting held on 17th March, 1971, as published in the July, 1971 issue of the Journal?' Agreed. WELCOME TO VISITORS The President: In welcoming you here, ladies and gentlemen, it gives me great pleasure to see so many of our office bearers-our important honorary office bearers-and particularly I would like to welcome our Honorary President, who is, as you know, the President of the Chamber of Mines, Mr John Shilling, and we are 'privileged also to have with us two of our Honorary Vice-Presidents. One in particular I am very pleased to welcome here, and that is Mr G. F. van L. Froneman, who is the Administrator of the Orange Free State, and we extend a very hearty welcome to him, because he is a long way from home. Then we have another Honorary Vice-President, Mr Tommy Gibbs, the Government Mining Engineer, and there are many distinguished guests, whose names I want to give you, so that you will know who they are. We are happy to have present: Mr R. C. J. Goode, Member of Council, I.M.M. London; Mr E. Boden, Manager, A.S. & T.S. of S.A.; Mr G. A. P. Louw, The President, The Associated Scientific and Technical Societies of S.A.; Mr I. R. G. Stephen, The President, The South African Institution of Electrical Engineers; Mr A. S. Robinson, The President, The South African Institution of Civil Engineers; Mr C. F. J. Schneider, The President, The Institute of Land Surveyors of the Transvaal; Mr R. J. Vermaak, The President, The South African Institute of Assayers & Analysts; Prof J. P. F. Sellschop, The President, Joint Council of S.A. Scientific Societies; Mr T. C. Watermeyer, The President, The Federation of Societies of Professional Engineers; Mr G. A. MacWhirther, The President, S.E.I.F.S.A.; Mr C. G. Hinds, The President, The Mine Managers Association; Mr J. F. Aldersley, The President, The South African Institute of Foundrymen; Dr H. O. Reisener, The President, The Institute of Welding; Dr F. P. A. Robinson, The President, S.A. Corrosion Council; Mr A. Bain, The Chairman, Witbank-Middelburg Branch; Prof D. J. Schoeman, The Chairman, The S.A. Institution of Chemical Engineers; Dr N. P. Finkelstein, The Chairman, The South African Chemical Institute (S.Tvl Section); Mr J. M. Meyer, Chairman, O.F.S. Branch. To all these guests and their ladies, I extend a most cordial welcome, and hope that you are going to enjoy the afternoon. MEMBERSHIP The President: 'I have pleasure in announcing that the names of the following candidates having been published in accordance with By-Law 5.2.2., Council has elected them to membership in the following grades: Fellows: Richard Lloyd Blandy, Byron J. Haley. Members: Brian Roger Borton, Frederick Brian Peacock, Peter John Stacey, Kenneth Niles Thanstrom. Students: Peter Boesley James, Kenneth Brian Perel. TRANSFERS From Member to Fellow: Cecil Eugene Fivaz, Peter Edwards. I wish to welcome the new members to the Institute and to congratulate those members who have been transferred to a higher grade. ANNUAL REPORT OF THE COUNCIL AND ACCOUNTS FOR THE YEAR ENDED 30TH JUNE 1971 THE PRESIDENT: We now come, ladies and gentlemen, to the annual report of the Council of the Institute-and our accounts for the year ended 30th June, 1971, and these documents are on your chairs; I know you probably haven't had time to study them yet, but I would like to highlight some of the things which I think are worth mentioning. I would ask you to take, first of all, ladies and gentlemen, the report of the Council as being read, and before moving adoption, however, there are certain salient points I wish to touch on, and I will leave the financial side of this entirely to our very competent and worthy, Honorary Treasurer, Mr Denis Maxwell.

Solvent loading: It is agreed that it is advantageous, as far as reagent cost is concerned, to operate at maximum solvent loading. However, owing to the fact that no reliable automatic solvent flow controller is yet available, control tends to be erratic with the result that there is risk of losing uranium. As can be seen from Mr Tunley's plot, the graph representing total ammonia cost is fairly flat between loadings of 3.5 and 4.0 gpl, while the graph representing uranium loss in the raffinate is steep. It is, therefore, safer to operate slightly below optimum loading in the absence of absolute control. Should the purity of the final product gain importance, operation at optimum loading or even slightly above, may be advantageous. Solvent losses and clarification of pregnant solution: The cost of replacing solvent, as indicated by Mr Carman, represents a high percentage of the overall Purlex costs. It is, therefore, agreed that, amongst other factors, efficient clarification of the pregnant solution and mixer-settler design have an important bearing on the cost. However, as at Blyvooruitzicht, the Buffelsfontein plant can tolerate a suspended solids content of 50 ppm without serious deterioration in efficiency. An increase in the suspended solids content from 15 to 50 ppm normally accounts for an increase of approximately 1 ppm in amine loss. Clarification of acidic pregnant solutions is much more difficult than that of alkaline solutions. Sand clarifiers are being used at Buffelsfontein with moderate success. Clarified solutions containing less than 20 ppm of solids in suspension can be produced, depending on the quantity and type of flocculant used in the filtration process, the thickness and the grading analysis of the sand beds and the quality of supervision. An on-stream duty of approximately 0.15 gallon per sq ft of surface area is obtained. Hydraulic handling of clarifier sand, while cleaning, may effect the efficiency of a clarifier, as size segregation of sand particles may lead to inefficient "pockets" in a sand bed. Control of the solution level in the sand clarifier also affects the quality of the effluent. Exposure of the sand bed invariably causes a cloudy effluent. Recycling of the final and initial solution discharged before and after skimming the clarifier sand improves the clarification efficiency significantly. Post-precipitation of silica and possibly sulphate salts, after clarification, may give a false impression of inefficient clarification. This phenomenon is more pronounced when treating hot supersaturated solutions. Ageing and cooling of the solution before clarification assist in stabilising the solution and reduce precipitation after clarification. The maintenance cost of sand clarifiers at Buffelsfontein is in the order of 0.35 cent per metric ton of solution treated. A redundant ion-exchange column has been tested successfully, either as a primary or a secondary clarifier, depending on the quality of the filtrate treated. The conventional pebble and sand beds in the column have been retained. Operation was enhanced by the addition of a two-foot deep resin bed as a filtering medium. The suspended solids content was reduced to less than 5 ppm. Mixer-settlers: Mixer-settler design is a controversial subject but, in view of the high costs of solvent compared to relatively low pumping costs, it is believed that mixing and pumping functions should be separated. Separate facilities for pumping and mixing not only give greater operational flexibility but also permit eventual changes in the flow circuit as technology develops. Scrub technique: Unlike the pilot plant, where the crud was carried away in the aqueous phase, crud is carried over with the organic phase in the full-scale plant. This carry-over of crud into the strip bank aggravates crud formation in the first strip stage. This phenomenon may result from a slight difference in settler design and/or a variation in the relative phase depths in the two plants. Bypassing of the scrub aqueous solution from the fourth to the second stage was introduced in order to avoid dilution and partial neutralisation of the acid added to the third stage, as it is known that a low pH value enhances the transfer of iron into the aqueous phase. The low aqueous flowrate in the third scrubbing stage was overcome by introducing an aqueous phase recycle. Gleaning: Mr MacDonald's contribution on "gleaning" was most informative. This process for recovering solvent from raffinate seems to show at least as much promise as coalescers and other methods employed to date. Similar tests conducted at Buffelsfontein pilot plant gave 60 per cent amine recovery from raffinate containing 10 ppm amine. However, the circuit was not operated at equilibrium and recovery may be expected to drop as the amine concentration of the "Gleaner" circuit increases.

By A. W. S. Schumann

A. W. S. Schumann The first reason for todays success lies in the very high standard which was set by our speakers today and I think that we owe a very special vote of thanks to Messrs P. M. C. Wilson, A. A. Oakes, M. J. Deats, A. C. Coetzee, R. E. Burnton and J. G. Ferguson. This, however, cannot account for the fact that there was a tremendous response even before anyone knew who was going to speak or what they were going to speak about. I suggest that the response was enthusiastic because people expected a more informal gathering with a greater opportunity for feedback, for an exchange of views and for questions and answers. The very popularity of the idea has defeated its own ends to some extent. This gathering is too big and many people must have been left with questions unanswered. Possibly one should arrange for a number of meeting places-there are many in the Chamber and the surrounding mining houses-and then have separate meetings in the morning with joint reporting session in the afternoon. Alternatively, we could have a symposium after the colloquium, using Webster's definition of a symposium as as 'a drinking party of merry fellows'. We all know that it is easy to discuss mining into the small hours. Obviously one cannot summarise todays papers and discussions, but I would like to make a few comments on technical matters and after that I would like to speak for a moment about what I see as major trends or tendencies which arise from discussions of this nature. Regarding the question of placing a village on a mine or away from a mine, I would mention that there is quite a body of literature on the adverse sociological effects on the development of children in any community in which the great majority of parents share the same occupation. It is something to be avoided if possible. We discussed the concept of the moving stockpile. If this is proved to be successful, it could save a considerable amount of money for Escom who could then avoid the construction of very large and costly staithes which are now regarded as parts of power stations. One does not see why both a moving stockpile and a staith should be necessary. By the way, speaking of Escom, I see that Mr Bob Scott is here; he has stopped pulling my leg by saying that we sell stone to Escom; perhaps someone has told him that if this is true, it must be the cheapest stone you can buy in this country. My next point relates to the use of epidiascope. It struck me again that this instrument which provides a visual image to accompany your message, can be very effective if it is properly used. These papers speak of a great deal of hard work which went into their preparation, and one really feels that the authors should have taken a few minutes more to prepare their drawings specifically for the epidiascope. There should be nothing on the drawing which is not necessary to illustrate a particular point and there should be nothing on the drawing which is not visible from the back of the hall. In particular, if one cannot read the figures along the sides, then one

Dr Hanekom I would like to thank Dr Hanekom for a very valuable -contribution; it summarizes careful statistical test work which was an omission from the original paper. Minor manipulative details, the application of a finer measuring graticule and the reproducibility of the procedure are described and the author accepts with gratitude this ready-made answer to the first and second points of Dr Finkelstein's comments. Mr Williamson The author is indebted to Mr Williamson for historical background to early experiments with captive bubbles and it is conceded that the original paper described another avenue to an old flotation goal. The field of application is restricted and we would not work with all the material illustrated by Klassen and Mokrousov.1 The particles in Figure 74(a) are far too large and irregular; material in Figures (b) through (d) could be employed for semi-quantitative work in spite of obvious disadvantages; Figure 74(f) represents ideal conditions. We have studied this mineral-collector association in the size range given and results were satisfactory. For larger pick-up values than that illustrated an induction period of more than 0.01 second was required (actually 90 seconds for a maximum bubble load at pH 7.2 with 25 g/t oleic acid!). REFERENCE 1. KLASSEN,V. I. ANDMOKROUSOV,V. A., 'An Introduction to the Theory of Flotation,' Butterworths, London, 128, Figure 74, (1963). Mr Slander and Mr Kooij Researchworkers in a practical mineral development laboratory may well ask what contribution yet another micro-technique can make to the problems of large scale ore flotation. The obvious answer is that there can be little or no direct translation of results obtained with a static system comprising air bubble and pure mineral concentrate to a dynamic multicomponent operation. However, for the rapid study of one mineral or variable some small scale apparatus must be employed. In the Anglo American Research Laboratory, the described semi-quantitative pick-up technique has filled a niche and requirement. For ourselves, the system represents a definite advance over the contact angle approach in that samples of minerals conventionally prepared in laboratory or plant may be used, which is a desired step nearer flotation. We bracket this pick-up apparatus with the Hallimond tube1,2 and evidence is now to hand that pick-up curves may often be related to flotation recoveries obtained from this tube with pure mineral samples and from the Fuerstenau glass cell3 with the total ore at low pulp densities. We would not wish to push the comparison of pick-up results past this stage for there is much published development work relating the Hallimond and Fuerstenau equipment to flotation on a larger scale. At the Anglo American Research Laboratory for example, a test in a Fuerstenau cell is often the final integrated step in laboratory investigations into the TORCO process. As these contributors rightly point out, pick-up tests can offer nothing towards the determination of optimum reagent dosage for large scale flotation. This is the function of bench-type flotation cells with up to 2 kg batches of ore. At this level, collector performance and consumption and physical properties of the froth are more comparable with large scale operation. The pick-up cell, Hallimond tube and Fuerstenau cell are essentially equipment from which the effect of control parameters can be derived. For the assessment of collector behaviour, the pick-up approach is rapid and economic and we offer here a pick-up contribution towards the flotation of Witwatersrand uraninite with anionic sulphonate collectors (Fig. 1). A large crystal of the local mineral was not available for contact angle measurements and owing to its low concentration in Witwatersrand gold-bearing reef, a small sample was laboriously obtained by tabling, superpanning and heavy medium separation. The pick-up values of a sized portion were determined by the described procedure in the presence of nine petroleum sulphonate collectors. Responses fell into two categories; "normal" curves (dotted lines) and a group of collectors which were very active at low concentration but showed depression of pick-up at the higher levels. The similarity in form of curves for collectors 1 and 3 was apparent and reagent 3 was found to be simply a diluent of 1. Otherwise, collector performance could not be related to either sulphonate content or molecular weight of the active constituent. However, reagent 2 gave the best performance in a bench flotation cell with regard to overall uranium distribution but, as predicted, reagent 1 yielded the best grades. Later pick-up tests with a similar class of collector have suggested that depression at high reagent concentrations may be due to "bubble armouring" whereby an hydrophilic air-liquid interface is produced which rejects mineral attachment. Prolonged investigations with the same mineral sample have not been required except for the study of uraninite which lasted for more than a year. Variation in pick-up response could not be detected over this period during which the mineral was stored under distilled water. The only visible change was a small efflorescence of lead sulphate, readily removed and decanted off after stirring with a glass rod in the sample tube. In its main role as a "trouble-shooter", the pick-up method is applied to fresh mineral samples taken from a flotation plant and soon discarded or from a current run in the ore-dressing laboratory. After production of a concentrate, a solvent cycle, as described in the original paper, may be required to remove heavy organics or other contaminants as when a sample has been taken from circuits containing detergents, grease or return flotation solutions. This solvent cycle will not remove chemisorbed collectors and we prefer not to work with minerals from flotation tailings or with pick-up cell discards. It has been thought preferable to answer the contributors' other apprehensions with regard to sample preparation, activators, depressants and the physical properties of reagents in another paper on offer to this Journal. It covers pick-up investigations relating to the flotation of nickeliferous pyrrhotite at two Anglo American Corporation Mines, which approach seemed the correct orientation for a reply to another contributor to this paper.

By D. G. Krige

D. G. Krige: The stimulating contributions from Messrs Cooke, Curtis, Martus, Johnston, Van Wyk, Van Rensburg and Wroth are appreciated by the author and provide clear evidence of the importance not only of the principle of assistance to gold mines, but also of its scope and form. In general all the arguments, raised by Messrs Cooke, Martus and Johnston in favour of extending the scope of the new assistance scheme can be accepted as sound.

By HORNICKA OSTRAVA

The Czechoslovak Scientific-Technical Institute has arranged the international conference 'Hornicka Ostrava' for four or five days at the end of October, 1969, as part of the preparation for the international 'Miners' Congress' on deep coalextraction in 1970. The keynote of the conference is 'New Techniques in coal mining,' and will be covered by five groups dealing mainly with: 1. New trends in the opening of mines, optimal parameters, projections, equipment, utilization of computers for economic analyses, etc. 2. Increase in efficiency of coal extraction in deep mines, technology, concentration, pressure control, optimization of material and coal transport. 3. Modernization of devices for increased safety and working conditions, new trends in explosives, new techniques for reducing environmental diseases. 4. Innovations in deep level coal extraction, mechanization, rationalization, automation, etc. 5. Other topics about modern techniques in mining. Address your enquiries to: Ing. Karel Kalas (Secretary of the Conference), Dum Techniky Cs YTS, Ostrava 1, Jesenkiho 7-9, Czechoslovakia.

By E. R. Leeman

INTRODUCTION The first references in the literature to the so-called C.S.I.R. 'doorstopper' strain cell were made in 1964 1,2,3,4. These were the climax of nearly two years of experimental and developmental work with the instruments and the operating techniques involved. Since that time numerous field measurements have been made with 'doorstoppers' and they have been developed to a stage where they are in commercial production and are being supplied to users throughout the world. No detailed description of the latest 'doorstopper' equipment has, however, appeared in print and this paper is written to bring the literature up to date in this respect. The triaxial strain cell was developed three years ago 5, 6, 7 and was first described at the First Congress of the International Society for Rock Mechanics in Lisbon in 1966 8. Since that date laboratory tests and-particularly important-field tests 7' 9 have demonstrated the validity and practicality of the instruments and techniques. These instruments have also been developed to a stage where they have been released for commercial production. In this form they are very different from those described two years ago and a full description of them at this stage seems to be justified. THE INFORMATION SUPPLIED BY 'DOORSTOPPER' AND TRIAXIAL STRAIN CELLS Both the 'doorstopper' and triaxial strain cells were designed specifically to determine the absolute stress in rock using an overcoring stress relieving technique. They were not designed to measure changes in stress. They may, however, be used for the latter purpose provided the glue which is used to stick the strain gauges to the rock is known to possess sufficiently stable strain-time characteristics as not to affect the accuracy of the results during the time the measurements are made. The 'doorstopper' was designed primarily to measure the major principal stress in situations where its direction and those of the two other principal stresses were either known or could be assumed. For example, at great depth in undisturbed virgin ground, it is reasonable to assume that the maximum principal stress is vertical and that the other two lie in a horizontal plane. The borehole in which the 'doorstopper' is installed would thus be drilled in the horizontal plane parallel to one of the minor principal stresses, the direction of which would either have to be assumed or determined as described later. The stresses measured would be the vertical major principal stress and the principal stress acting in a horizontal plane normal to the borehole axis. It is possible, as will be shown later, to obtained the complete state of stress using 'doorstoppers'. In this event, it is necessary to make measurements in three boreholes drilled in any three known directions to each other and relative to a system of three orthogonal axes. The triaxial strain cell was designed to determine the complete state of stress in a single borehole drilled in any direction in any stress field. At first glance it would appear to supersede the 'doorstopper'. This is not necessarily true since there are

By C. Fiore

The annual general meeting of the branch was held in the Witbank Golf Club on Thursday, 28th August, 1969 at 5:00 p.m.

Nov 1, 1969

By A. F. Lee

Dr E. C. Hanekom (Visitor): Two years ago we used the bubble pick-up technique as developed at Anglo American Corporation for screening candidate flotation agents prepared in our laboratories. It was obviously necessary to know to what extent the technique gave reproducible results and a statistical evaluation of the test method was made using a mineral sample provided by Anglo American Corporation. A bubble of 2.0 mm diameter was used and the measuring graticule was divided into squares (units) with sides of 5.0 mm at 25 X magnification. The mineral granules had diameters from 53 to 74 mm. The bubble was loaded by depressing it against a layer of mineral grains to give an arbitrary but constant compression of 0.4-0.5 mm. It was rolled five times through 2 mm horizontal displacement. The bubble was then lifted, its load compacted by gently stirring and the projected area measured in squares (units) on the grid. Each substance was tested at 5 ppm concentration in distilled water at five points in the pH range 3-6. With these standardized conditions of operation the reproducibility of the test was studied. Four materials were tested and at each pH value the procedure of forming a bubble, picking up granules and counting the units was repeated ten times. From these results the standard deviation for a single count at any pH value was calculated for each material (Table 1. The coefficient of variation of approximately 3 per cent was relatively low and it was decided that further refinements in the counting procedure would be of doubtful benefit with the present apparatus. Similar experiments were carried out with 12 sample materials. In this case the bubble formation and pick-up were repeated four times at each pH value, the mean value was calculated for each set of four observations and final pick-up values for the materials were obtained by selecting the highest mean count. Each test was completely repeated a number of times as indicated in Table 2 and the standard deviation calculated. The results allowed the selection of the more promising materials and the rejection of poor performers at the 95 per cent confidence level. All tests were carried out by one operator. Some evidence was obtained that the method is sensitive to operator bias. In conclusion it can be said that from all the results analyzed it would appear that once operation conditions are fixed (viz operator, pH control) the precision of the method is adequate. The author wishes to thank Messrs African Explosives and Chemical Industries Limited for permission to publish these data. G. S. Stander and C. J. Kooij (Visitors): We must congratulate Mr Lee for this most interesting and useful development of the pick-up experimental method. It certainly provides research workers with an instrument which makes flotation more of a science and less of an art. We visited Mr Lee at his laboratory where he gave us a demonstration of the bubble pick-up instrument and technique as developed and improved by him.

Jan 12, 1969

By A. M. Starfield

Discussion J. de V. Lambrechts (Fellow): The author's paper is a brilliant follow-up of an earlier paper by Starfield and Dickson.1 I have no quarrel with Dr Starfield's computerization of a complex problem, but I do not believe that he is getting quite the right answers from his programme. My impression is that his predictions about wet bulb temperature increases in very deep mines are over optimistic; in other words, that it will be hotter than Dr Starfield predicts. This is putting my views in a nutshell. This is not the occasion on which to indulge in lengthy argument about the original paper by Starfield and Dickson, but the present paper is, after all, based directly on that earlier paper and if the one fails, the other cannot succeed. I did level certain criticisms at the first paper and cannot say that the authors' replies were very convincing. I do not think it is a sin to admit that I belong to the old school which believes in thorough field experimentation and practical trials and no amount of mathematical manipulation or physical theorizing, no matter how excellent, can make up for inadequate practical confirmation. This, as I see it, is still the crux of the matter. The original paper by Starfield and Dickson still rests on somewhat scanty practical evidence and I would be much happier if Dr Starfield's computer programme, based on the Starfield-Dickson model, had been checked against a large mass of observations in the practical mining situation. This is what both Wiles2 and myself3 had tried to do previously. What we lacked in mathematics and/or computer aids, I think the present paper by Dr Starfield lacks in practical substantiation. This is no condemnation of the author's paper which, taken by itself, is excellent but I think the final stage is still lacking, namely the bringing together of theory and empiricism in a manner acceptable to all. This may be wishful thinking on my part but I hope, within the next year or so, to come up with a modified Starfield-Dickson model in such a way that the computer answers will agree in the majority of cases with the few hundred field observations which are already on record. It might be a case of applying the proverbial 'Cook's Law' to the Starfield-Dickson model! REFERENCES 1. STARFIELD, A. M., and DICKSON,A. J. 'A study of heat transfer and moisture pick-up in mine airways.' J. S. Air. Inst. Min. Metall., 68, (5), 1967. 2. WILES, G. G. 'Wet bulb temperature gradients in horizontal airways.' J. S. Air. Inst. Min. Metall., 59, (7), 1959, p. 339. 3. LAMBRECHTS, J. DE V. 'Prediction of wet bulb temperature gradients in mine airways.' J. S. Air. Inst. Min. Metall., 67, (11), 1967, p. 595. R. Hemp (Visitor): Dr Starfield's paper has very effectively rounded off one particular aspect of the general problem of heat flow in mines. The ease with which this computer programme can be used to calculate temperature increases in horizontal airways must lead to its wider use in ventilation planning and, in developing this rapid method, Dr Starfield has indeed rendered a valuable service to the mining industry. One could consider further instances of heat flow in airways in which the availability of a rapid computer method would be desirable, e.g. the flow of air down a shaft, where there is an increase in temperature due to adiabatic compression, as well as an increase in virgin rock temperature as the depth increases. However, this particular case would not present any new problems and would merely require an extension of the exisitng work. I should like in this contribution to talk about an aspect of environmental control in mines which, I think, will become more important in the future. It is well known that wet bulb temperatures are subject to fluctuations underground. In some instances, particularly in stopes, the fluctuations, both with time and position, can be considerable. The theoretical work which has been carried out on temperature increases has been aimed at the prediction of mean temperatures, and no account has been taken of fluctuations around this mean. It is questionable whether this approach will, on its own, be sufficient, particularly when temperature increases in stopes are considered. The fluctuations in air temperatures underground arise from two causes. The first of these would be the fluctuations in surface conditions, and here one could list random, diurnal and seasonal fluctuations. The second cause is the multitude of things which vary in a mine and here one could list variations in air flow quantity, sources of evaporation, heat transfer from pump and compressed air columns and, particularly in the stope, variations in air flow patterns. Fluctuations arising from surface temperature variations should be amenable to calculation, and here one envisages figures relating the decay of temperature variation with distance to factors such as air flow quantity. Fluctuations resulting from changes in the mine are perhaps more difficult to tackle theoretically and the best approach could well be to analyse underground observations. In this connection, there is a good case to be made for the increased use of statistical methods in the analysis of underground temperature measurements, and it might be of value to look at current air-conditioning practice. When carrying out cooling load calculations for a particular location it is customary to use design wet bulb

Jan 11, 1969

By J. H. Viljoen, A. Heyns, J. F. Morrison, C. H. Wyndham, N. B. Strydom

Discussion M. J. Martinson* (Associate Member): This is another interesting paper in the series which the Human Sciences Laboratory is publishing on the physiological cost of the more strenuous manual tasks performed underground. As a mining engineer-albeit presently an academic one -this contributor is more concerned with the practical significance of the study in the field of stope cleaning than with physiological niceties, but before discussing some of the practical issues raised by the paper perhaps the authors would care to reply to lay comment on three physiological matters. Firstly, in two places the authors make statements to the effect that the maximum oxygen intake of 'average' (sic) Bantu mine workers is 3.0 l/min. What do the authors mean by 'average'? Maximum oxygen intake in the individual is probably some function of inter alia age, physical fitness, state of health, ambient partial pressure of oxygen, diurnal and/or seasonal changes in metabolism, and possibly cumulative exposure to dust and other occupational contaminants. Since maximum oxygen intake is apparently such an important quality in manual labourers we ought to have a clear picture of its distribution in the entire labour force, and also know precisely how the individual's quantum is affected by factors such as those mentioned above. Secondly, arising out of the same statements and ignoring the reservation implicit in the previous paragraph, is it correct to say that the 'average' Bantu mine worker has a maximum intake of 3.0 l/min? In the August 1966 issue of the South African Journal of Science four of the present co-authors reported the mean maximum oxygen intake of a random sample of 338 Bantu recruits of different tribes to be 2.63 l/min with a standard deviation of 0.542 l/min, and these figures were repeated in a paper published in the November 1968 issue of this Journal. Admittedly the mean maximum oxygen intake of the six subjects used in the present study was 3.05 l/min, but since they were selected on the grounds of previous shovelling experience they can hardly be regarded as being 'average' recruits. Thirdly, it seems to me that the authors are perhaps a little uncritical in their acceptance of the '50 per cent of maximum oxygen intake' level of activity for endurance work. I am not unmindful of the evidence upon which this level is based, but I suggest that the evidence is at best tenuous and that the whole topic of fatigue and 'optimum' levels of productivity under conditions of high muscular activity and considerable psychological stress is a lot more complicated than would appear from the authors' ready acceptance of the 50 per cent level. Under present working conditions the question is largely rhetorical because it would need an army of supervisors to ensure that all members of a typical gang work at 'optimum' levels throughout the shift, but if individual performance should ever become more readily determinable the issue might become crucial to management and malayisha alike. So much for physiological matters; now a few random thoughts on some wider, practical issues raised by the paper. In the paragraph headed Material the statement is made that 'The mixture (of fine, medium-sized and bigger rocks) was considered to be representative of an average underground rock sample'. Elsewhere in the report the authors specifically draw attention to the effect of particle size on the oxygen consumption of men shovelling graded material, and intuitively one imagines that the size distribution of the material being shovelled would be highly significant in a study such as this. Presumably the authors were forced to assess size distribution subjectively because of the lack of objective data on the subject, and possibly the necessity for this assumption highlights how little we know quantitatively about any of the variables in stope cleaning systems. In this connection it may be remarked that although the present study was largely designed to show how dip and stoping width effect energy expenditure in shovelling, no data are given-nor, probably, are they available-to show how the industry's stoping tonnage is distributed in terms of these two factors. It seems to me that if the mining industry seriously hopes to improve stope productivity it must start by measuring the basic variables in existing systems. Elaborating somewhat on the previous paragraph, there also seems to be a tremendous gap between the present study-and its predecessors-and the prospect of gainfully using the information so carefully garnered by the authors. It would be reassuring to hear from the authors that their work on the physiological aspects is being integrated in some master plan for increasing productivity of manual labour. This I fear is not the case, and if productivity has changed at all in recent years I would hazard a guess that the change has been brought about by random factors rather than by systematic engineering design. The Human Sciences Laboratory has worked extensively in stopes in the past, and possibly the time has come when it should widen the ambit of its operations to include engineering-type analyses of cleaning systems, with particular emphasis on the economics of rapid cleaning by manual labour. Perhaps experiments could be conducted with stakhanovite-type gangs consisting of men selected for their high maximum oxygen intake and superior motivation; these men might be offered special incentives in the shape of extra pay

Jan 10, 1969

By C. H. Wyndham, N. B. Strydom

Discussion C. G. Hinds (Member): The climatic chamber at Kloof Gold Mine was built to the design referred to in the paper presented by Professor Wyndham and Dr Strydom. As no suitable underground facilities existed for the acclimatization of labourers at the rate required, mining operations would have been severely restricted had there not been a climatic room timeously built on surface. The climatic room can accommodate 120 labourers per 4-!-hour shift. The standards laid down by the Human Sciences Laboratory were and still are being adhered to as strictly as possible. Through a shortage of European personnel the recommended degree of white supervision could not be adhered to and the use of Bantu assistants with a minimum education of Standard 8 was introduced. This is considered to have met with sufficient success to warrant acceptance as policy. The climatic room was brought into operation on the 1st September, 1967, i.e. two months after stoping at Kloof was commenced. The figures quoted herein therefore pertain to the period between that date and the 31st March, 1969, when the main shaft and the main surface fans were commissioned. The areas worked were sited between 5016 ft and 6 365 ft below surface. Virgin rock temperatures ranged between 92°F and 99°F. Stopes are cleaned by hand lashing. During the 20 months of stoping prior to the commissioning of the main shaft, mining operations at Kloof were effected through the Harvie Watt shaft of the Libanon Gold Mining Company Limited. In this period 10 452 Bantu labourers were acclimatized in the following categories: Experienced workers 7 326 New to the industry 2 086 Re-acclimatized 1 040 It will be noted that during the period, labourers 'new to the industry' averaged 22 per cent of the intake. This percentage fluctuated from time to time and reached 42 per cent. (The present percentage is 60.) The greatest number acclimatized during anyone month occurred in February, 1969 when 1234 Bantu were treated. The greatest number of Bantu acclimatized during any 8-day period was 411. As many as four X 41/2 hour shifts were run in 24 hours. The number of men found to be heat intolerant over the period of 20 months was 33 out of a total 9 412 Bantu. The following tabulation illustrates the history of the build-up of manpower and the climatic conditions under which they worked: The average dip of the reef is 35°. The number of heat stroke and heat exhaustion cases during the period of 20 months under most difficult ventilation conditions was Nil.

Jan 10, 1969

By E. Pothas, P. R. Jochens, L. B. McRae, D. D. Howat

Discussion H. J. S. Kriek*: During 1956 Iscor became interested in the acid smelting in a blast furnace of siliceous ores containing 15 per cent silica. This interest was as a result of the successful acid smelting in Germany at a CaO/SiO2-ratio of between 0.4 and 0.8. Acid smelting can become an economic proposition due to a reduction in slag volume accompanied by a lowering of the coke rate. Paschke and Hannel1 found that a free flowing slag (i.e. viscosities of less than two poise) would not be obtained at C/S-ratios lower than 0.82 and that the addition of 3 to 4 per cent alkalies became necessary. In titanium bearing slags it was established that maximum fluidity is obtained with 10 per cent TiO2. Acid smelting resulted in increased sulphur pickup with decreasing O/S-ratio and desulphurization with soda-ash outside the blast furnace was necessary. At Iscor Mr B. B. Segal established that, for slags with (CaO+MgO)/SiO2-ratios of 0.5, 0.6 and 0.7, the addition of alkalies and TiO2 resulted in a substantial decrease in melting point and also of viscosity. At a (CaO+ MgO)/ ratio of 0.53 the addition of 10 per cent TiO2 is as effective as the addition of 3.75 per cent alkalies. This means that a blast furnace can be run without the addition of alkalies and consequently the detrimental effect of alkalies on the lining can be avoided; also that fluid slags with 5 to 15 per cent MgO and 8 to 10 per cent Al203 can be obtained with 10 per cent TiO present. (CaO+ MgO)/SiO2-ratios smaller than 0.5 the addition of titania resulted in a decrease in melting point and also in an increase in viscosity. In ceramic crucibles the slag did not become more viscous with time but thickened appreciably when held in graphite crucibles. No titanium carbonitride could be detected and only Ti20a was found. Prof Howat also did not obtain the TiC-TiN solid solution and therefore it appears that the thickening of the slag is not solely due to the formation of the titanium carbonitrides. This is in agreement with the work of Michailov and Belyakova2 who found that TiO2 decreased the viscosity of acid slags and increased the viscosity of basic slags; that titanium sesquioxide (Ti203) and titanium monoxide (TiO) increased the viscosity of both acid and basic slags and that titanium carbide, which is found in larger amounts in basic slags than in acid slags, considerably increases the viscosity of both acid and basic slags. Patzak et al3 describes the following oxygen deficient titanias, viz Ti2O, TiO, Ti203, Ti3O5, Ti4O7, Ti5O9, Ti6O11- Ti10O9,Ti7O13,Ti8O15 and Ti9O17. Until about 1959 a fair amount of data was available on the viscosity of normal slags to which titania was added2, 4, 5, 6, 7, 8, 9. At Iscor an electromagnetic viscometer based on the viscometer developed by Bockris10 was built. Progress was very slow as our main attention was centred around problems associated with the operation of the Rotor process. During 1962 and 1963 two papers containing fundamental work on the viscosities of slags in the systems CaO-SiO2-TiO211 and CaO-SiO2-TiO2 at 0 per cent A1203, 10 per cent Al203 and 20 per cent Al20312 appeared. The work of Ross and Ohno obviated the necessity of any further work by Iscor on the viscosity of blast furnace slags containing titania. Ross gave the following optimum compositions for blast furnace slags containing titania. Titania % Lime % Silica % Lime/Silica-ratio 20 30-42 38-50 0.6-1.10 30 27-36 34-43 0.63-1.06 40 24-32 28-36 0.66-1.14 The optimum temperature is 1,500°C because at this temperature the largest workable composition range exists. Higher temperatures result in narrower optimum composition ranges and in increased rates of thickening. The factors affecting the rate of thickening were also investigated by Ross and Ohno, as the main problem with slags containing titania is not finding compositions which are fluid at operating temperatures but the thickening due to a lack of oxygen. With these slags when there is insufficient oxygen to co-ordinate each silicon atom with its own individual shell of four oxygen ions, some sharing of oxygen ions must occur, resulting in the formation of large silicate ions and a rise in viscosity. The formation of oxygen deficient titania can be prevented as long as iron oxide or manganese oxide remain in the liquid slag. Iwase13 and also Viens, Camp bell and Rogers14 obtained in the experimental smelting of ilmenite ores fluid slags with more than 3 to 5 per cent manganese oxide and ferrous oxide in the slags. It appears that titaniferous ores can be smelted in blast furnaces under the following conditions: (1) Acid slag operation, i.e. at C/S-ratios from O.6-1.0. (2) Dolomite can be used as a flux as MgO is a network modifier. (3) The Al2O3-content must be less than 15 per cent as alumina promotes thickening of the slag and decreases the optimum composition range. (4) The hearth temperature must not exceed 1,500°C. According to Ross a definite amount of energy is required for each ton of iron produced and in order to increase production an increase in heat

Jan 6, 1969

By R. R. M. Cousens, W. S. Garrett

SYNOPSIS The paper gives a background to the water hazards of the so-called West Wits Line and refers to some geological aspects of the Wonderfontein Valley in which the West Driefontein mine property lies. It describes the precautions taken against flooding during shaft sinking, development and stoping; these include cover hole drilling with cementation, provision for water storage in worked-out areas, installation of water tight bulkhead doors and pumping capacity well in excess of the mine's normal requirements. The plan of campaign adopted after the inrush took place is explained and the reorganization and expansion of pumping arrangements are discussed. When it became clear that the only means of saving the mine was to construct plugs to isolate the inrush while still maintaining all emergency pumping measures, the work was put in hand with the utmost dispatch. In considering the design of the plugs a number of major factors had to be taken into account. The most vital was the need for speed which was complicated by difficulty of access to the plug sites because of the huge volumes of water flowing in the drives. How this and other factors affecting plug design and construction, leakage problems, sequence of operations and final closure of the valves, were overcome is detailed in the paper. INTRODUCTION The inrush of water into the West Driefontein mine, the biggest gold producer in the world, which began on 26th October, 1968 and continued unabated for 23 days before the flood was stemmed, was unprecedented in the history of mining. Also unprecedented were the tremendous efforts and organization which made it possible to bring the situation under control and so save the mine from being lost for a matter of years. BACKGROUND Before going into details of how the situation was brought under control it is appropriate to consider the background to the event. The water problem on the so-called West Wits Line has been recognized for many years. As far back as 1912 the Pullinger Shaft, situated in the centre of what is now the Venterspost mine property, had to be abandoned because there were no means then available of coping with the large quantity of water encountered. To appreciate what vast quantities of water were involved, it helps to refer to some geological aspects of the Wonderfontein valley in which the West Driefontein property lies. The narrow tabular Witwatersrand gold-bearing beds, which dip at about 25 degrees towards the south, are overlain by a surface layer of dolomite which reaches a maximum thickness of some 4,000 ft at West Driefontein. A number of vertical syenitic dykes intersect the valley more or less at right angles to its course (Fig. 1). These dykes are about 150 ft thick on average and are spaced at about eight-mile intervals in the West Driefontein area. They penetrate both the dolomite and the Witwatersrand beds and as they are impervious to water they divide the valley into a series of large underground water compartments.

Jan 5, 1969

By R. R. M. Cousens, W. S. Garrett

Discussion: L. T. Campbell Pitt: When Mr Garrett and I discussed the design and construction of plugs and bulk head doors to resist great water pressures in mines about ten years ago, we visualized an emeregeny when there would be no time for site preparation. We could not however, anticipate a problem of such magnitude and urgency as the one described in the paper under discussion. The paper very clearly describes the flooding and the measures taken to save the mine. The complete flooding would have been a national disaster. The account of the combat between man and water reads like the account of a military operation. It was an endeavour which achieved success because there was a master plan, tenacity and, in fact, heroism. The paper is of necessity a long one. In my contribution I will mention some historical facts and factors that I believe have relevance to the flooding and the saving of the mine. It is of great satisfaction to me to have been one of the two initiators of a new approach to underground plug design and construction. The investigation was taken over by a Chamber of Mines Committee which combined our early work with further research and the resulting data had a significant bearing on the West Drie emergency plug designs. The final plug length was correctly based on the requirements to obviate leakage. A plug to have structural strength only would have been very much shorter but so ineffective as to have become a waste of labour, material, and in consequence, wasted the limited time available to prevent flooding of the whole mine. The figure of 75 lb per sq in. sheer stress provided a very ample margin, in my opinion, but under the very difficult site conditions and the limitations of plug material suitable for the very unusual construction, justified such a figure. There must have been considerable temptation to reduce the safety margin due to the urgency to complete and commission the plugs, but it is much to the credit of the directors of this operation that nothing less than a safe design was permitted. In about 1945 Blyvooruitzicht Gold Mining Company, which was then mining in to the West Driefontein area to assist in early development, decided to protect themselves from a possible inrush of water from the latter mine. A bulk head door was accordingly designed and installed on 2 level between 1945 and 1948 at 4,218 ft below surface. Since West Driefontein paid for it, Gold Fields consulting mechanical engineers collaborated with their colleagues of Rand Mines who completed their design after a few modifications were agreed. This door was never used to prevent an inrush into Blyvooruitzicht but the exercise was very useful to Gold Fields when the high pressure doors initially installed on 10 and 12 levels between Nos. 2 and 3 shafts to enable isolation of one section of the mine from another were designed. At that time the capacity of West Driefontein to cope with an inrush of water was comparatively small because the storage in stoped areas and the installed pump capacity were much less than in later years. These doors together with others of the same design were later moved to the lower levels served by sub vertical shafts. In addition, doors to prevent flood water entering the pump chambers on 18 level were installed at each end of that level, and designed for a head of 360 feet so that water would have to reach 16 level before the pumps were flooded. All these doors were called upon to hold water and it is reported that only a small plug in one of the doors and a 4 in. dia flanged pipe joint failed. The lesson is that in these cases the smallest details warrant as much attention as the major ones. That these were the only equipment failures when so much depended on equipment, is very much to the credit of design engineers, both of the Gold Fields head office and companies who supplied both mechanical and electrical equipment. Until I had read the paper I wondered why the policy of isolation by bulk head doors was not continued as the development moved eastward towards the Bank Compartment and particularly when it pierced the Bank dyke. The reasons have been given in the paper. It was unreasonable to expect provisions to be made to counter the remote contingency that the large installed pumping capacity and mine capacity to store water would be exceeded during a peak inrush. When pumps were being considered for West Driefontein, it was decided that the Gold Fields standard underground main pumps at that time of 1 million gallons per 24 hours would be too small. Three alternatives were considered. 1. Very large units that would require underground sub assembly and dismantling of both pump and motor. 2. The largest pump and motor that could be conveyed as a complete pump or motor from surface to the pump chamber. 3. Pumps of higher speed to reduce dimensions and weight. It was decided to adopt alternate 2 because of avoidance of transporting pumps in parts and assembling underground. Alternative 3 was not favoured because high speeds would have introduced problems of more exact rotor balancing. I believe the decision was correct and in the crisis enabled comparative easy transport of pumps and motors to surface for overhaul and, if required, transfer to other pump stations. The standard pump thus became a 10 stage 1,500 rpm pump with a capacity of approximately 2 million gallons per 24 hours against a static head that varied

Jan 3, 1969

By D. Rankin, H. J. Stucke, W. K. B. Loftus

Discussion: H. J. Wright (Member): The authors are to be congratulated on presenting such a comprehensive and lucid account of bringing a large diamond mine into efficient production from the grass roots. When one considers the vast amount of planning and hard work that had to be put into the project the authors have done a praiseworthy job in condensing the story, so as to present all the important technical aspects, the logic behind the major decisions and the future thinking on the operations, in a concise and readable form. This was the first experience of De Beers personnel in starting a large diamond mine as an opencast project and nor do they employ a large planning staff, most of the planning being carried out by the executive staff themselves. It is operating most efficiently on a large tonnage and it is neatly and very well laid out. In fact, with its model townships, both for European and Coloured employees, tarred roads, recreation club and its grass golf course, it is like an oasis in the arid climate of the North-Western Cape. The key operation of opencast mining is, of course, the planning of the geometry of the pit. This has been adequately dealt with by the authors. It is of interest to consider some of the consequences arising from complete lack of open pit planning, based on actual operating mines. I refer of course to the present, four producing mines in Kimberley. The Kimberlite pipes were all discovered in a short span of time from the year 1870, onwards. They soon became a hive of activity with diggers industriously working their claims at varying rates. The country rock comprises from the surface downwards, some 90 ft of decomposed dolerite followed by approximately 260 ft of soft Karroo shales, which in turn is underlain by a hard competent Ventersdorp lava, known locally as melaphyre. The average water table was just below the dolerite at a depth of approximately 90 to 100 ft. Only the blue ground was mined leaving the waste rock sides almost vertical, and as the depth increased beyond 100 ft, large falls of sidewall started to take place. The shale was unstable and the falls covered

Jan 3, 1969

By M. Berjak, D. M. Hawkins, V. E. Marting, R. B. MacGillivray

Contributions to discussion V. E. Marting (Associate Member): The authors of this paper should be congratulated in having pioneered the application of the Sharapov predictor to local data, and in having made a valuable contribution to South African experience in the quantitative treatment of geological information using a digital computer. Although the existing set of three programmes has been described as a 'suite' it does not represent an integrated system in the accepted sense since the communication between runs is made in the form of punched card files to enable the mining engineer to exercise discretion and control over the data entered at each phase. Although the authors are justifiably proud of the successful results obtained through the application of this method, I feel that it is my duty to comment that the computer aspect of the work has not reached full development and is deserving of the financial support needed to increase its widest usefulness. At the present moment a routine to detect 'missing information' has not been commissioned. Due to the lack of documentation the use of the programmes is restricted since the format for the submission of data is known only to the authors. It is very unlikely that a profitable return on the development costs will be enjoyed before the programme is fully documented and made available to wider use. My criticism of the present incomplete state of the work is being voiced purely with the intent of focussing on the significance of this project and the fact that it is deserving of development on an industry basis. It should be noted that at the economical cost of R4.00 per sq ft of computer print-out, a mining engineer can hardly afford to miss the opportunity of using the programmes on other (e.g. metalliferous) mineral deposits of tabular formation. Since the publication of this paper I have had an opportunity to see the results of the programmes developed by Drs M. D. G. Salamon and F. Deist for designing bord and pillar workings. The basic data for these latter programmes are extracted from iso-line plans drawn conventionally by geologists, a task that could be considerably expedited through the use of Mr Hawkins' programmes. It appears that the efforts of these respective teams are complementary and it is in the interests of the industry that these individual components should eventually form part of an integrated system. E. Margo* (Member): This paper is a good example of how and where the backroom scientist-in this case the mathematical statistician-can aid the mining engineer to develop not only more accurate means for dealing with valuation problems but to develop a tool that can also cope with changed geological specifications.

Jan 1, 1969