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By James S. Scott

Effluents discharged from gold mills, and more recently from heap leaching operations, pose significant hazards to the environment if not properly managed prior to discharge. This paper describes the nature of wastewaters which must be treated, the treatment processes being applied in Canada and the capabilities of these methods in removing cyanide and metals. As a result of past practice the most common method of reducing cyanide in gold mill effluents remains that of natural degradation in tailings ponds. This method has proven capable of producing high quality effluents where large ponds and sufficiently long water retention times can be provided, but at many mines this has not been the case. Since 1981, as a result of these limitations there has been a strong move in Canada to chemically-based effluent treatment plants, particularly as additional processes became available. Canadian gold mills have become leaders in the use of chemical processes to remove cyanide, in terms of the variety of such processes in operation (five in total), the number of plant installations and the number of processes developed in this country. The hydrogen peroxide and Inco SO2-air processes, in that order, are being applied most widely. Alkaline chlorination, the initial process to be used fairly extensively has been almost totally replaced by the two methods mentioned above. Since a number of effective methods are now available for the treatment of gold mill effluents, the selection of a process for application at a specific mill requires careful consideration of:

Jan 1, 1987

By Harold Rabling

THE following notes are taken from results obtained on a standard 25-ft. Hancock jig 1 tested during regular operation in the Bonne Terre mill of the St. Joseph Lead Co. The object of the tests was to determine the conditions for most effective 'work, and the nature of products that could be made. The material treated was sized between 9mm. and 2-mm. round-hole screens, and was constant throughout the experiments, Three products were made, concentrates from the first three hutches, middlings from the fourth and fifth, and tailing from the sixth. The middlings were crushed in one pair of Allis-Chalmers Style B 30 by 14-in. rolls, and elevated with the original ore to the 2-mm. screens, oversize of which formed the jig feed. While this arrangement was very simple, it gave rise to undesirable crowding of the circuit. Table 1 shows tonnage and assay values of the jig feed and products at the beginning of the experiments, and brings out the excessively large amount of middlings, comprising 63.3 per cent. of the total material fed to the jig. The rolls, being heavily overloaded, did very little crushing, a screen analysis of the product showing that only 27 per cent., or 128 tons, was crushed fine enough to pass to the tables. The remaining 347 tons returned to the jig with 400 tons of original ore. At times, this overcrowding became much worse, and the tonnage in circuit grew to enormous proportions. It became imperative

Jan 8, 1917

By C. E. Roessler

Uranium and its decay series are associated with phosphate deposits of marine origin. Uranium concentrations are on the order of 100 to 200 ppm in matrix from the Bone Valley formation of Central Florida and about 25% this value in matrix mined from the Hawthorne formation in North Florida. Occupational radiation exposure due to this source has been under study in Florida since the mid 1970's (Windham, Partridge, Horton 1976; Roessler and Prince 1978). This report is based on findings of an industry-wide scoping survey conducted by the University of Florida during March 1976 through May 1977 and subsequent studies in selected individual facilities. PHOSPHATE MINING OPERATIONS Mining and beneficiation operations are diagramed in Figure 1. At the excavation site, electric powered draglines remove the overburden and expose the phosphate matrix (ore) which is then slurried and pumped to the beneficiation site. At the washer stage of the beneficiation plant, the matrix is subjected to washing and screening. A coarse product, "pebble" is produced and sent to storage piles while the fine fraction is passed through a series of cyclones to remove the clays. The generated clay suspension is sent to settling ponds while the remaining sand-sized fraction is sent to the flotation plant. At the flotation plant, phosphate particles are recovered from the fine sand particles of similar size through several stages of treatment with reagents, agitation and separation of floating and sinking fractions. The high grade concentrate produced is transferred to storage while the sand tailings are discarded. [ ]

Jan 1, 1981

By James Norman

SPODUMENE is a lithium-bearing pyroxene, and is an important source of lithium compounds. Because of its high alumina and lithia content it might be a desirable constituent of glass batches. The use of spodu-mene in ceramics will undoubtedly increase, as it has been shown that a mixture of the proper proportions of this mineral with soda feldspar and potash feldspar has a much lower maturing temperature upon firing than has any one of the three components.1 The mineral spodumene seldom occurs as pure mineral, but ordinarily with feldspars, mica and quartz. All spodumene-bearing rocks tested contained these minerals as gangue. Small amounts of iron and manga-nese oxides also were present, usually as stains on the silicate grains. Mica associated with the spodumene carried a considerable amount of lithia. This was discovered in flotation tailings, which, although virtually free of spodumene grains when examined microscopically, showed 0.4 to 0.5 per cent Li20 on chemical analysis. The lithia content of micas picked from the North Carolina spodumene rock was variable. Some samples contained little lithia, others more than 3 per cent. In 1938, under the supervision of Oliver C. Ralston, ? the Bureau of Mines, in cooperation with the Tennessee Mineral Products Corporation, undertook to study the beneficiation of North Carolina spodumene by froth flotation. The Bureau had previously published investigations on the beneficiation of spodumene.2,3 The methods described in the second publication' did not give good results with badly weathered spodumene rock, so that more thorough investigation of the subject was warranted.

Jan 1, 1940

The Mount Lyell Company's concentrating mill is situated about a quarter of a mile westward of the portal of the North Lyell tunnel, which serves the North Lyell, Royal Tharsis and Crown Lyell mines, being on the 1100-ft. level of the first named. The Comstock mine is connected with the tram system at this level by five and a half miles of steam tramway, two-foot gauge being standard throughout.The plant is well placed with regard to all essential services and has an excellent water-supply from a series of three dams on the West Queen river.MILLING ORESThe mill feed is a mixture of the outputs from the foUr mines mentioned, each of which has its characteristic ore, but, generally speaking, about half of the ore consists of siliceous schists and the other half of hard grey quartzite. The schisty ores are easy to crush, but, on grinding, while the schist slimes readily, the quartz granules are hard, resulting in the handling of heavy loads on the secondary classifiers. The quartzite ores are both hard to crush and difficultto grind.The principal sulphide materials are bornite, chalcopyrite, and pyrite, chalcopyrite being the main copper-bearing mineral.The crystallization is coarse in the schisty ores, but it has been shown that even the 200 mesh (Tyler) material contains included sulphide particles. At the present time the sizing of the final tailing averages 80% through 200 mesh.

Jan 1, 1932

By George A. Packard

WHEN, in April, 1892, the writer began experimenting with the cyanide-process, it had already proved a success in the treatment of tailings, but had not become an important factor as a primary method of ore-treatment. The Livingston, Colorado, mill was running a few small agitators; the Mercur mill, in Utah, was said to be a success, but was " closed down to increase capacity; " and Mr. A. B. Paul, in Shasta county, California, was running ore'through a stamp-mill into leachingvats, using a weak cyanide solution in the battery. Mills had been, or were being, erected in Arizona; on the Comstock lode; and in South Dakota. Some of these never started, and several were shut down or remodelled for other processes. Altogether, the outlook was anything but favorable. Many of these mills were planned to make use of agits tion, and a cyanide solution containing at least 1 per cent. of potassium cyanide was employed. Since then the improvement of the process, both chemically and mechanically, has placed it among the recognized successful methods of oretreatment. Within the past year I have visited a large number of cyanide-mills, and have collected considerable data, of which I publish enough herewith to give an idea of the development of the process and of the methods followed in the principal mills. I would say, however, that nearly all of these plants are experimenting, most of them employing experienced chemists,

Jan 1, 1897

By J. L. Gaigher

A simple mineralogical regression model is described, which was devised in order to help the flotation plant cope with rapidly changing headgrades and feed mineralogy. Minerals in the feed were determined by means of quantitative X-ray diffraction and compared with plant performance parameters and flotation reagent consumption rates, using regression analysis. A mineralogical regression model for predicting recovery served to motivate flotation plant personnel and also set standards whereby flotation performance could be measured. The reagent consumption model indicated in which direction any or all of five flotation consumption rates should be altered. All practically measurable parameters, even those showing poor or no correlation, were included in the model, provision being made for the continuous update of a batch of data of 30 lines. Compared with the method of multiple linear regressions, the r2 is poorer at 0,41 (versus 0,50), but the mean standard error of estimate for both methods is identical at ± 6,6 units. The regression model is simpler than the method of multiple linear regressions, and although the accuracy of prediction is not high (41%), the mineralogical regression model proved to be successful in practice. Flotation efficiency, as measured by the mean ratio of head to tails grade, for the 9 month period after the mineralogical regression model was implemented on an hourly basis, improved from 2,8 to 3,2, which is equivalent to an improvement in average flotation recovery of 5% relative.

Jan 1, 1987

By J. Herbst, H. B. Treweek, V. R. Degner, G. Poling

Current mineral industry trends toward handling the high tonnages of lower grade ore and tailings retreat lead to a continuing need for large volume flotation cells. The benefits which accompany the larger cell flotation circuit include reduced capital cost, less mill floor space, reduced operational complexity, and less maintenance, provided the metallurgical performance and solids suspension capability of the cell are not compromised. This study reviews development considerations pertinent to large mechanical air induction type flotation cells from the standpoint of cell internal hydrodynamic characteristics and metallurgical performance. While the importance of process related factors such as chemical usage, pH control, solids size and loading, and circuit arrangement cannot be denied, this study will emphasize the relationship of cell size (volume), mechanism geometry (impeller diameter), and mechanism operating conditions (rotor speed and submergence), to air transfer, liquid circulation, liquid "rise velocity" and mechanism power draw. Specific hydraulic performance maps for large capacity 300 and 500 cu ft flotation cells will be reviewed in terms of the relationship between the operational variables and the air and liquid flow conditions within the cell. Metallurgical results, obtained under plant operating conditions for 300 and 500 cu ft cells, are included to illustrate the metallurgical performance similarity of flotation cells which follow hydrodynamic scale-up criteria set forth in this investigation.

Jan 1, 1976

By W. L. Zeigler

The mill is a departure from gravity concentration and has gained a reputation for the low initial cost of erection, extreme simplicity and the low cost of milling on the refractory character of the ore which is treated. The structure is of heavy timbers with board walls and corrugated roofing. The floors are concrete and the general layout is of the ground floor only. All the machines are driven by individual motors and the transmission usually found in such a plant is entirely absent. (Fig. 2.) ORE The ore from the Galena mine is sheared Wallace quartzite replaced with quartz, siderite, pyrite, sphalerite, gray copper (tetrahedrite) and very fine-grained argentiferous galena. Much of the silver is undoubtedly associated with the gray copper, the amount, however, apparently varies with different localities in the vein. An average assay of the ore is 6.0 per cent. lead, 1.3 per cent. zinc, 0.3 per cent. copper, 17.5 per cent. iron and 4.5 oz. silver. The concentrates produced are of an average 53 per cent. lead, 40 oz. silver, 8 per cent. zinc and 11 per cent. iron, while the tailings are 0.5 per cent. lead, 0.4 oz. silver. With the present market for zinc no attempt is made to produce a zinc concentrate.

Jan 1, 1927

By Plyley WF

Kidston Gold Mine is a 3.5 million tonnes per year operation located 280 kilometres west of Townsville, Queensland. Kidston's ore body consists of 44.3 million tonnes graded at 1.79 grams per tonne gold and 2.22 grams per tonne silver. Full production commenced in March, 1985. During the remainder of 1985, Kidston milled 3.77 million tonnes of ore and produced 213,614 ounces of gold.The ore is comprised of granodiorite, polymictic and volcanic breccias. The degree of oxidation of the rock influences the milling capacity and therefore determines the cut-off grade; ore is categorised as oxide, transition and fresh. The milling rates are 15,400, 12,000 and 9,600 tonnes per day for oxide, transition and fresh ore respectively.The ore is treated in a conventional cyanidation/ carbon-in-pulp plant (CIP). The plant consists of a primary crusher, semi-autogenous grinding (SAG) and ball mills, leach circuit and carbon-in-pulp circuit. Carbon stripping and dore smelting are done on site. The major difference between the Kidston plant and other Australian CIP plants is that a cyclone separation and washing circuit rejects half of the dry tonnage as a sand tail prior to CIP treatment. This difference allows the CIP circuit at Ki.dston to be smaller and makes the CIP less sensitive to the grind circuit product size.

Jan 1, 1986

By Otto F. Pfordte

The great advance of the last twenty years in the specialization, delicacy and efficiency of ore-concentrating apparatus calls for a finer system of testing both the qualities of ores and the operations of machines in this department. In amalgamation, lixiviation and smelting, the daily employment of physical and chemical tests has been recognized as necessary, and the cost of the necessary laboratories and employees is cheerfully paid. But while much progress has been made in concentration (due, in a large degree, to the study of the subject in our technical schools), it is still too common in actual practice to follow the loose old method of roughly estimating, once for all, that a given ore " concentrates x tons into one," and resting on this assumption so long as that ore continues to be treated, while assays of the tailings from concentration are made from time to time, without exact determination of their quantity. The first step toward accuracy of control must be a more frequent and precise measurement of the quantity of concentrates from a given operation. This should be expressed in pounds per unit of one ton; and the variations shown in frequent determinations will furnish a constant check upon the character of the ore and the performance of the apparatus. But this is only a part of what should be constantly and systematically done to insure the best work upon a material which must vary continually, even though it come from one mine only. If, in any mill, appropriate laboratory-tests were made at least once a week, the labor and expense involved would be amply repaid by increased knowledge and control of the current operations. A small testing-laboratory, provided with a few gold-pans;

Jan 1, 1902

By F. H. McDowell

The electro-magnetic separator of Mr. G. Conkling, Glens Falls, N. Y., described by Mr. Birkinbine* in a paper at the New York meeting of the Institute, in February, 1889, has been in successful operation at the Tilly Foster 'mine for a considerable period, under the disadvantages which seem to be the inevitable attendants of pioneer enterprises in new technical fields. The results of a run of several months, which the Superintendent, Mr. S. P. Tomkins, has taken pains to collate, are here presented by his courteous permission, and that of Mr. E. S. Moffat, the energetic and skilful General Manager of the Lackawanna Iron and Coal Company, to whom our profession is already indebted for examples of bold and ingenious practice in mine-engineering. It is believed that these figures will assist those who are planning similar operations to decide intelligently what they may, and, more important still, what they may not, expect to accomplish. The local difficulties in the way of this experiment at the Tilly Foster mine were many and great. Only about two-thirds of the waste-dump, resulting from the assorting of previous years, is ore; and this varies in content of iron from 20 to 28 per cent. The remainder is rock, some (from horses in the vein) showing particles of adhering mineral, some (from the foot- and hanging-walls) entirely barren. The expense of overhauling the dump and removing this rock is by no means the least of the numerous items which must be charged against the concentration. Associated with the lean ore are hornblende and other minerals, carrying iron in the non-magnetic state, which must necessarily be washed away with the tailings. The waste-dump was originally located without reference to future treatment, and now can only be handled by means of a locomotive,

Jan 1, 1891

By Gary F. Rajala, Bruce R. Greenwood

INTRODUCTION The Empire Mine is an iron ore mining, concentrating and pelletizing facility operated by The Cleveland-Cliffs Iron Company in the Upper Peninsula of Michigan. Owners of the mine are Inland Steel, LTV Steel, Wheeling-Pittsburgh Steel and Cleveland-Cliff s. The orebody is a fine-grained, low grade iron formation with magnetite as the primary constituent. Also present are iron carbonates, iron silicates, chert, hematite, goethite and quartz. The original six concentrating lines were started in late 1963. Successive expansions of ten lines in 1967, five lines in 1974 and 3 lines in 1979 have resulted in Empire reaching a present annual capacity of 8.0 million long tons of pellets. The Empire flowsheet was developed over several years of laboratory bench testing, pilot plant testing and testing of a full-scale primary autogenous grinding circuit. Since the original flowsheet was established, modifications have been made based on pilot plant and full-scale plant tests. The present flowsheet for three sections of the concentrator consists of one stage of crushing, two stages of closed circuit autogenous grinding, two stages of magnetic separation, one stage of hydraulic concentration and two stages of flotation (rougher and scavenger). The flowsheet for the fourth section of the concentrator (Empire 4) is the same as the first three sections, except a crusher has been added to crush excess pebbles. The Empire 4 flowsheet appears in the at- tached figure. PRIMARY GRINDING CIRCUIT FLOWSHEET Empire has a total of 24 individual grinding lines. However, at the present time, Line 1 is not being used for processing crude ore. The primary mill on Line 1 is now grinding flux stone for fluxed pellet production and the secondary mill is being used as a ball mill to grind rougher flotation tailings prior to a scavenger flotation circuit.

Jan 1, 1992

By James Barr

WHEN phosphate mining operations first commenced in Tennessee the loss of both high- and low-grade material was large, because of the crude hand methods employed. Practically all rock smaller than 2 in. was thrown back into the pits and covered with the overburden of subsequent openings. With the advent of mechanical washers and other mining machinery, together with the abolishing of the contract system, the waste has been steadily diminished, until at present only the finest sand is being rejected; and even this is not lost, for in the majority of cases the water, slime, and sand from the washers are run into large settling ponds where practically all of the phosphate-bearing material is caught and stored for future use, or until it is profitable to ship lower grades. The latest equipment used to recover the fine and lower-grade phosphate sand includes Dorr thickeners, settling tanks, Dorr classifiers and settling ponds served by locomotive cranes with clam-shell buckets, making it possible to save material averaging as fine as 200 mesh. To attempt saving finer material would necessitate a different type of drier, which might be patterned after the multiple-deck roasters. The present tailing runs from 15 to 25 per cent. of available phosphate, which, however, as mentioned before, does not represent a total loss to possible future operations. The demand of the consumer is the greatest factor in regulating the recovery at present. In the mining of blue rock, which does not require washing treatment, the main waste material is the capping of low-grade kidney forma-tion, generally used for backfilling of the rooms. By reason of the introduction of improved methods in the early stages and the re-mining of territory skimmed over at the start, at the present time the net result of all operations is a comparatively small loss.

Jan 2, 1916

` LAWRENCE ADDICKS, New York, N. Y.-We have really two competitive processes for the production of electrolytic zinc. In the process used at Great Falls, the ore is roasted and the calcines leached, leaving a residue which has to be smelted; in the other method, the ore is burned, the zinc and lead fumed off, and the condensed fumes are leached. I do not think that we can yet come to any decision as to which is the better way to attack the problem. Of course if we roast first and then leach, various substances remain in the calcines which may consume sulphuric acid; whereas if we do the burning first, we obtain simply the zinc and lead oxides, with considerable sulphate, in the fume, and there is no shortage of acid. It also depends partly on the complexity of the ore. The one that Mr. Hall is treating and the ore they are treating at Great Falls are not quite alike, but I think they are sufficiently similar to make a comparison of the two methods, working in competition, on different parts of the country, very interesting. C. E. SCHWARTZ, Miami, Okla.-Mr. Hall states that the Joplin conditions are not suitable for the manufacture of electrolytic zinc. When we speak of Joplin ore in general, we consider, of course, a 60-per cent: concentrate with comparatively low iron, but if a low-grade concentrate could be made from tailings and so-called chats, then the problem would be quite different. A Joplin ore is comparatively simple to treat electrolytically, and I believe the electrolytic process would open a field for the treatment of such ores as would not be ordinarily smeltable; that is,

Jan 1, 1918

By E. Mcl. Tittmann

Deep-seated human instincts urge us to positively mark the passage of time. We celebrate the passage of each year. Years give way to decades, and decades to half centuries and centuries. At all these mileposts we are inclined to pause, take a look around and catch our breaths before embarking on the next leg. We are at such a point now and this meeting I am sure will give direction to the road ahead into the decade of the seventies. In looking around at this point, 1973, it is obvious to all of us that the environment in which we must mine, mill, smelt and refine lead and zinc has radically changed. These changes have, of course, occurred over a long period of time. They have, however, all been in the same general direction. They have all had the same motivation. The period of greatest contrast is probably the fifteen years just before World War II and the fifteen years following the end of that war. I thought that I might review briefly with you what seems to me to be four of the more important changes that have developed since World War II that profoundly affect our activities in the lead and zinc industry. I mentioned before that all these changes have been in the same direction; that is, they have all tended to restrict our freedom of action in pursuing our objective of producing lead and zinc on a competitive basis at a profit. Consequently, in discus- sing them you may think that I am merely viewing with alarm. This is not the intent. This is merely supposed to be an objective analysis. I think it appropriate before discussing these trends in de- tail to examine the position of lead and zinc today.

Jan 1, 1970

By R J. Nicholson, J M. F Clout

As part of an overall strategy to make significant new technological advances for the sustainability of the mineral processing and related process industries CSIRO Minerals has developed a new dry separation device. It represents a technology breakthrough and has the potential to radically change the way in which dry particle size classification is achieved in the minerals industry. Conventional dry separation techniques have limited applications and often suffer from low productivity, high operating costs, low separation efficiency, the need for dust control and a lack of flexibility. In the absence of efficient dry techniques, wet processing is often used. The use of wet processing either requires the introduction of water into a separation process, which causes environmental concerns through the need for water supply and tailings dams or the use of æunfriendlyÆ chemicals for dense media separations that have considerable impact on handling and disposal. Wet processing also increases operating costs due to plant instability, materials handling, thickening, filtration and product drying. The new CSIRO technology exploits a combination of natural and applied forces to achieve specific particle motions that produce size separation. In doing so, many of the traditional shortcomings of dry separation have been overcome. The CSIRO Dry Particle Size Separator is ideally suited for dry size separation applications within the size range 1 mm to 30 ¦mfor moderate moisture levels (2 per cent), have low unit throughput, produce a fixed cut size and are therefore impractical where very high unit capacity and low operating costs are required. This paper presents a review of dry separation devices together with details of the new Dry Particle Size Separator and typical results for a number of materials, including iron ore and quarry dust.

Jan 1, 2002

The paper gives details of the performance of a 6-ft. by 6-ft. grate discharge ball mill which was converted into a rod mill of the overflow type and then into a grate discharge rod mill. Proposals for further alterations are outlined.INTRODUCTIONThe mill of King Island Scheelite (1947) Limited recovers scheelite by gravity methods.Crushed ore at the rate of 34 tons per hour is divided between two mills. (6 ft. by 6 ft. and 8 ft. by 6 ft. respectively), the mill discharges being pumped to a common distributor which feeds seven Hummer Screens. Each screen has three 4-ft. by 2-ft. Ton-cap sections and repulping compartments are provided. The screen oversize is returned to the mills whilst the screen undersize is finished grinding section product. A third (6 ft. by 6 ft.) mill regrinding the coarser table tailings also discharges into the common Hummer Screen Distributor.The screen undersize is pumped to a bowl classifier which separates at 200 mesh. The bowl classifier sands are treated on tables whilst the bowl classifier overflow, after thickening, is treated in the slime plant by flotation. The flotation concentrate is further concentrated on vanners and slime tables.A recovery of 80 per cent. is possible by tabling whereas only 50 per cent can be recovered in the slime section. The object of comminution is, therefore, to reduce all the new feed to less than a certain mesh (22 mesh Tyler) with the production of a minimum quantity of minus 200 mesh material. cheelite is, a very friable heavy mineral (Sp.Gr. 6'0) and is therefore susceptible to overgrinding. For the evaluation of grinding performance it is hence advisable...

Jan 1, 1955

By Francis S. Kendorski

INTRODUCTION Nuclear waste has been generated since the time of the Curie's and continues to be generated worldwide at an ever-increasing rate as the world will (and should) rely more and more upon fission (and hopefully fusion) for power. Also, the needs for strategic weapons will also generate so-called defense wastes. The largest class of high level nuclear waste (HLW) is spent fuel, the tail end of the nuclear reactor fuel cycle. At present, spent fuel is stored at the reactors producing it or at a few accepting places for away-from-reactor storage. All such storage facilities, as presently constructed, will begin to run out of usable storage space during the next decade or so. Congress enacted and President Reagan signed into law in early 1983 the National Nuclear Waste Policy Act of 1982. This law provides for the permanent isolation of nuclear waste in deep geologic repositories. The Act sets 1998 as the date at which the first repository should be operational. This date is extremely ambitious, and the Department of Energy (DOE), which is charged with the responsibility of constructing and operating repositories, admits that this date will slip by at least a few years. The Act also provides for licensing of repositories by the Nuclear Regulatory Commission (NRC) in a fashion very similar to nuclear power generating reactors. This requirement imposes a very special rigor on the analyses, investigations, designs, construction, and operation of repositories so as to assure the protection of the public and the environment from radioactive contamination. In addition the states and affected Indian Tribes must be consulted, especially concerning siting, and, needless to say, no one wants a HLW repository in their back yard.

Jan 1, 1984

By A. Gillis, J. Keyes

"This is a prelirr1inary report on automatic control at East Malartic Gold Mines.After installation of the cyclones in the East Malartic grinding circuit, the interest was aroused in the automatic control system for the circuit. Investigations were made as to the possibilities of such an installation. This resulted in a preliminary survey of the circuit by Milltronics Ltd and an analysis of the circuit by Mines Branch personnel. A study of the investigation and analysis indicated that a cost reduction and improved efficiency could be expected by the installation of an automatic control system in the grinding circuit. It was also indicated tl1at such installation should pay for itself within a year. This was the point that sold management on such installation. The need of three operators was expected to be eliminated .The installation was to consist of a central control panel, microphoncs on each mill, regulation of feed rate, pneumatic valves for addition of solt1tions, density control of the cyclone overflow, which is a grinding circuit product, rod mill feed box plugging detector and a conveyer weightometer for tonnage indication and control.As well as the elimination of one man per sl1ift, other benefits which were to be received but unable to estimate were: 1) More tonnage per clay by operation of the mills at the proper capacities. 2) Lower grinding costs. Less under-loading should result in less steel wear and lower pumping cost. In the condition of overloading, a coarser grind is obtained resulting in greater impeller and liner wear. The overloading and underloading showed up on the n1icropl1oncs. 3) An improvement in the recovery as a result of less over-load. The tailings loss from a coarser grind are not as frequent."

Jan 1, 1965