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By Walter von der Linden

INTRODUCTION There are three essential components for a modern longwall mining system: the coal cutting/loading machine, the self-advancing supports and the longwall conveying system. The latter consists of a face conveyor, a stageloader and a belt tall section arrangement (Figure 1). The longwall face conveyor is an integral component of the face equipment and has a threefold purpose: 1) continuous coal haulage. 2) track for cutter/loader . 3) reference/anchor rail for roof support advance. Further requirements for the longwall conveyor are horizontal flexibility for conveyor advance without interruption of haulage and cutter/loader tramming and vertical flexibility to cope with seam undulations.

Jan 1, 1981

In 1972 the Marra Mamba Iron Formation in the Lower Proterozoic Hamersley Iron Province of Western Australia was recognised as a stratigraphic horizon with significant potential for high grade iron ore reserves. Within the Newman area the Marra Mamba Iron Formation has been subdivided into four strati- graphic units: Lower BIF member, Middle BIF and shale member, Upper BIF member and Upper shale member - and a standard natural gamma log reference section compiled for the entire formation. The recognition of the Upper shale member as an additional unit contrasts to the previously proposed stratigraphy. Enrichment of the Marra Mamba Iron Formation to ore grade (0)8 percent calcined Fe) in the Newman area is restricted to the Upper BIF member and to a lesser extent to the Middle BIF and shale member. Typical ore from Orebody 29 is friable interbedded martite-ochreous goethite, signifi aptly different both physically and chemically from the Brockman Iron Formation ores of the Tom Price-Whaleback type. Large scale customer acceptance trials (in excess of 4 million tonnes) have shown that Marra Mamba ore crushed to fines and blended with Mt. Whaleback fines in a ratio of up to 15:85 forms an acceptable sinter feed.

Jan 1, 1981

By Alan R. Gadberry

Good mine planning is critical to the success of a safe, cost effective mine. Many factors contribute to the overall success of an open pit mining operation: proper pit location and development, good drilling and blasting practice, and well conceived short range and long range mining plans.

Jan 1, 1981

Mr. Chairman, Distinguished Guests, Ladies and Gentlemen. I was delighted to be invited, and to accept the invitation, to welcome to this specialist meeting on an intriguing subject, the local participants and - most importantly - the international contributors. Technical o analysis of blast furnace operations has for many years formed the basis of intense research throughout the iron and steel industry, world wide, because the efficiency of blast furnace operation estab- lishes the foundation for all subsequent processes in a steel plant. So much depends on the constant, efficient performance of iron melting units in relation to costs, quality, regularity of operations and reli- able delivery performance to our customers.

Jan 1, 1981

By A. H. PH. D Wilson

Compared to the rocks encountered in, say, metalliferous mining or tunnel drivage in igneous and metamorphic rocks, the strata associated with the coal measures are relatively soft. If excavations are made at depths greater than one or two hundred metres, zones of fractured or yielded rock develop around the openings. The stress which can be taken by yielded rock at the boundary of an opening is low, but because of the friction present between the fragments or particles, the sustainable stress rises rapidly with distance from the opening until sufficient confinement has been developed to prevent failure of the rock in the first place. Thereafter the laws of elasticity will apply and increasing distance will reduce the stress field until conditions consistent with the cover load are reached. For rocks which maintain their elastic condition right to the edge of an opening, finite element analysis and similar techniques have proved successful in determining the distribution of stress around the opening. This approach, however, is not successful when a yield zone is introduced between the elastic zone and the opening. Nevertheless, some estimate is required if help is to be given in the siting of roadways and in the determination of adequate pillar sizes. This matter is of particular importance in the planning of new mines especially if retreat faces are involved, as extensive roadway drivage must be planned and executed before experience of the effects of face extraction is gained. In 1977, the author published [I] a method based on 'stress balance'. The total aggregate downward load over a large area will remain that of the cover load even after part of the coal seam has been removed. Any rise in stress in ribsides and pillars must be offset by an equivalent stress reduction across roadways and other areas of extraction, and vice versa. Knowledge of one can lead to an estimate of the other, provided the general form of the stress distribution can be postulated. It is the object of this paper to revue the method in more detail and to give further examples of its successful application.

Jan 1, 1981

By Yutaka Mochida

OUTLINE Beginning in 1946, a careful geological survey was made of the Tsugaru Straits. In 1964, excavation of the exploratory tunnels were started to observe the characteristics of faults and rock formation directly, and to study construction methods of excavation and sealing water inflow. In March 1971, the investigation stages were over, and in April 1971, the construction project was authorized from the government. The geological features in the straits may be divided into three broad classes. One third consists of andesite on Honshu side. The other two thirds consists of tuff. The middle of the straits consists of tuff, so-called Kuromatsunai-layers. On Hokkaido side, it consists of tuff, called as Kunnui-layers. There are nine noteworthy faults on the route. Prior to the excavation of the main tunnel, the horizontal pilot tunnels and service tunnels are made. Exploratory borings are taken to confirm the geological conditions and the presence of gushing water. Groutings for sealing water inflow are carried out when necessary. In geologically bad conditioned sections such as disintegrated or fractured zones, special methods are applied. In geologically good zones, the tunnels are bored through using the same methods as for a mountain tunnel driving. The pilot tunnel being constructed at the sea bottom is under direct control of the Japan Railway Construction Public Corporation. The service tunnel and the main tunnel are being excavated by contractors.

Jan 1, 1981

By William A. Hockings

Methods for the measurement of liberation of finely ground iron ores are described. The general method consists of separating a sample of the ground ore into a number of size fractions, followed by measuring the liberation of each fraction. The size separation has been done using electro-formed screens of 20-, 10-and 5-micrometer apertures, and also by air elutriation. Liberation (i.e., the distribution of compositions of individual particles) has been measured by optical microscopy. A computer pro-gram was used to adjust the raw count data so that it agrees with the chemical analyses of the size fractions. Various ways of presenting the results are given, and Some applications of the results to flotation processing are discussed.

Jan 1, 1981

By M. K. Abel

"The Strathcona Mine is located in Levack Township on the North Range ofthe Sudbury Basin. The main orebody consists of pyrrhotite-pentlandite-chalcopyrite mineralization associated with an embayment of the footwall of the Sudbury Irruptive. A zone of massive chalcopyrite-cubanite-pentlandite mineralization occurs in a complex vein system in a series of mixed gneisses and breccias in the footwall. Vein widths vary from a few millimetres to 6 metres. In December 1979, this zone was estimated to contain 976,000 metric tonnes, with a grade of 9.31% Cu and 0.46% Ni. The vein system has a general strike of N 68°W at an angle of 60 degrees to the average strike of the footwall contact of the Irruptive. The dip of the vein system is 45 degrees sw. The strike and dip of individual veins in the system coincide with the directions that would result if a compressive stress had acted in a direction normal to the footwall of the Sudbury Irruptive, producing a conjugate set of fractures. The strike of the Irruptive at the Strathcona Mine is parallel to the long axis of the basin.IntroductionThe Strathcona Mine, owned and operated by Falconbridge Nickel Mines Limited, is located in Levack Township on the northwest rim of the Sudbury basin (Fig. I) . It is one of a group of nickel-copper sulphide ore deposits present along an 8.2-km section of the North Range of the Sudbury Irruptive. In the mines of the North Range, subunits of the Irruptive, going from hanging wall to footwall , include felsic norite, dark norite and dark norite breccia. The footwall complex at Strathcona includes the late granite breccia, which hosts a large proportion of the nickel-copper ores, and a compositionally variable group of brecciated and migmatized rocks, herein referred to as the feldspathic gneiss complex. It is within this latter group of rocks that the vein system of the Copper Zone occurs."

Jan 1, 1981

By Louis A. Panek

INTRODUCTION AND SUMMARY With respect to the geomechanics aspects, the primary technical objectives in mining by an undercut-cave method are to achieve a controlled, sustained caving of the mineral body and to remove the fragmented ore with a minimum of dilution from surrounding unmineralized rock, while maintaining stability or control of the rock structure around the access openings. The structurally ideal ore body probably does not exist. If the rock mass is so weak as to cave on a short span, it may tend to be too sticky for easy drawing and handling, or create special problems in regard to support of the access openings. If the ore is strong, caving may be difficult and the caved fragments may be of such a large size as to require special equipment for transferring the ore from the caved zone. Given enough time and money to generate data and conduct trials, engineering and ingenuity can devise an appropriate combination of mine layout, sequence of extraction, and mechanical equipment to achieve a technically successful caving extraction operation to meet the foregoing requirements in many types of deposits. The large capital investment involved, however, reduces the freedom to make major changes once the mine development is well under- way, and the penalties for failure to accurately anticipate operating conditions militate against selecting any but the most obvious candidates for mining by an undercut-cave method. The demonstrated capability to extract large, deep, economically marginal deposits by this low-cost, high-volume method of mining provides an incentive to develop a rationale for predicting the cavability and stability characteristics of a deposit prior to mining, so that the undercut-cave method may be extended to a much wider range of mineral deposit characteristics. The ultimate goal is to establish as explicitly as possible the quantitative interrelationships between the measured rock-mass characteristics, the caving span, the size distribution of caved ore fragments, and the sizes and locations of stable access openings. Lacking an understanding of these relation- ships, a designer may readily change some factor in the wrong direction (e.g., excessively reduce the distance between the extraction level and the undercut to increase the convenience of operations for the undercutting crew, increasing the frequency of repairs to the extraction-level support system) or create unnecessary problems elsewhere in the system by introducing a design change that can achieve only minimal improvement in the factor of direct interest (e.g., unnecessarily complicate the ore-transfer system by changing the orientations of the openings, with- out succeeding in the objective of improving the ground support conditions). Although successful predesign is the prime objective, subsequent modifications in mine lay- out and sequence of extraction operations are inevitable. In developing the modified solution, systematic experimentation based on an understanding of the underlying structural relationships, coupled with monitoring measurements of selected diagnostic structural-behavior parameters, can achieve an acceptable solution in a minimum number of steps, which is far superior to the typical operational trial-and-error approach, in view of the cost of implementing each successive change. Since a drill-core sample of ore rock from a successful undercut-cave operation may exhibit a uniaxial crushing strength in excess of 100 MPa, the caving of such a rock mass is now commonly believed to be ascribable to the presence of discontinuities such as joints or fractures throughout the ore body. An essential part of the present investigation was therefore to characterize the natural discontinuities at each of the test sites by measuring their attitudes and spacings. The term "fracture" is used herein in a general sense to include any planar discontinuity without implication as to its suggested mode of origin. Most, but not all, of the fractures are properly termed joints. As a point of departure we may consider the possibility that the rock mass is transected by three families of joints, each family possessing a distinct orientation, such that parallelepipeds of intact rock are delineated by the jointing. Even if cementation is absent between adjoining parallelepipeds, the undercutting of the rock mass will not necessarily initiate sustained caving--owing to the all-around confinement, an arch may tend to stabilize over the undercut unless prevented from doing so by the failures of key blocks of intact rock. Thus, although the jointing can be assumed to weaken the rock mass, creating preferred directions of

Jan 1, 1981

By Torres S. Ricardo, Encina M. Víctor, Segura O. Claudio

INTRODUCTION This paper is a conceptual resume of engineering studies that have been evaluated by the Mine Department related to production planning over the next twenty-five years. These studies have been required principally because: a) The Rio Blanco Mine, property of CODELCO- CHILE, Andina Division, faces an operational problem due to heavy snowfall in winter which enters the surface subsidence area of the cave; later melts and filters into the broken ore creating a product that is difficult to transfer through the orepasses. The phenomenon is especially serious during the spring and has been increasing over the years. This element that complicates extraction will be controlled with the procedures proposed or others that may be determined in the future. b) The mine will increase production in 1985 by 40% . LOCATION The deposit is located in the central region of the Andes Cordillera, at an altitude of approximately 3,600 m above sea level, 50 km to the northeast of Santiago. The main camp is Saladillo (at 1,300 meters above sea level) located 30 km from the mine. Due to its location, the mine is subject to heavy snow storms during the year. On many occasions the road from Saladillo to the mine is closed because of storms and avalanches. For this reason the company has self-sufficient civil facilities near the mine to cover the needs and welfare of the working force. These housing facilities are connected to the mine by underground passages. For the same weather related reason the primary and secondary crushers, and the entire concentrator complex are located underground. GEOLOGIC CHARACTERISTICS Geology Based on the geologic model developed for the Rio Blanco deposit and comparing it with standard models pertaining to copper porphyries, it can be concluded that this deposit is part of the apical sector of a mega deposit and therefore the projections of the Xio Blanco cop per mine are very favorable, both in lateral area and depth. In order to confirm and evaluate these geo- logic features, an extensive drilling program was started at the Rio Blanco deposit. Diamond drill holes were drilled to depth as well as in the areas named "La Americana", "Don Luis" and "Elguin". The area explored by the drill program is 3 km long in a north-south direction and 1.5 km wide in an east-west direction. In some sections the drilling in depth reached the 2,300 m elevation mark. The lowermost level in the mine is at elevation 3,070 m. These new studies have permitted the recognition of the following lithologic units: Andesite Andesite Breccia Granodiorites, Rio Blanco and Cascada Granodiorite Breccia Quartzmonzonite Porphyry Monzodiorites and Latites Tourmaline Breccia Dacite Pipe and Rhyolite marginal ring. The mineralogical study permits the definition of three stages of hypogene and hydroterma1 alteration of importance with a time-altered sequence. These stages correspond to a late hydrothermal stage. The kinds of alteration that have been recognized are: Potasic Alteration Propylitic Alteration

Jan 1, 1981

By Karl H. Gehring

According to the increasing importance of boom type roadheaders in tunnelling essential problems occuring in the application of these machines are dealt with. Based on a chart including the various stages for decicion finding the excavation method of underground cavities, both the technical limits and economic criteria are discussed in order to determine the range of applicability of roadheaders. Giving examples-the numerous developments in the fields of pick and cutting technology, performance data of boom type roadheaders in hard rock, safety technology and increase in the machines' degree of utilization, research and development of VA is illustrated. Two topic operation examples show partial aspects of the various tendencies of development.

Jan 1, 1981

By Alfonso W. Ovalle

INTRODUCTION The block-caving mining method is one of the most economical ways existing today to extract ore from nature. It is undoubtedly the least costly of the underground systems and moreover it compares favorably to open-pit mining. In addition to its low cost characteristic, the block-caving mining method permits a great extraction volume as can be exemplified by El Teniente's output capacity of over 22 million tons of ore per year. These characteristics of block-caving give assurance that, even though it is a very old mining system, it will be a good answer to the many future large massive low ore grade deposits to come and a1 so to rep1 ace open-pi t mining as such operations get deeper with higher stripping ratios. These are the reasons that induce further study of block-caving. In this paper block-caving at El Teniente is analyzed with due consideration to the major problems that this mine will encounter. Emphasis is given to the factors influencing the mining of soft ore which accounts for most of the ore to be extracted in the next ten years. A separate paper is devoted to a major modification introduced to block-caving hard primary ore at El Teniente by means of LHD equipment extraction. DESCRIPTION OF THE PRESENT MINING Ore Production and Flow Over 62.000 tons per day of ore are extracted 359 days per year at the El Teniente Mine. The Teniente 8 train haulage level at 1.983 m takes 29.000 tons daily to the Col6n Concentrator. Two trains make the 8 km haul with 16 cars per train. Car capacity is 90 tons. The Teniente 5 train haulage level 298 m above takes 33.000 t per day to the old Sewell Concentrator. Ore is produced from three production areas in two different levels. The so called North Mine produces 42.000 tons daily from the Teniente 4 level 65 m above the Teniente 5 haulage level. The Central Mine produces 1.000 tons from the first hard rock block caved at the Te- niente 4 level. The combined 43.000 tons per day production is sent through inclined 2 m dia- meter raises to the Teniente 5 haul age level , where 33.000 tons are taken 2.5 km to the Sewell Concentrator and 9.000 tons are hauled 0.5 km to a 5 m diameter vertical ore pass which sends ore to Teniente 8. The haulage is by 7 trains, 15 gabble-bottom cars per train. Car capacity is 20 tons. The so called South Mine produces 19.000 tons per day from the Teniente 1 level 645 m above the Teniente 8 haulage level. Ore is transferred by 2 m diameter inclined raises to the Ten 1 Retram level 22 m below, where 5 trains with 20 Sandfor Day cars per train, 8 tons capacity per car, haul the ore 1.5 km to two vertical 5 m diameter ore passes which connect to Ten 8. Starting in 1982 the Central Mine will in- crease production to 6.000 tons per day. An in- dependent loop haulage system will start operat- ing at the Teniente 5 level to feed a new under- ground 1.372 mm (54 in.) gyratory crusher which will send minus 152 mm ore (9 in.) to Teniente 8. In two or three years time the North Mine will only feed the 33.000 tons to the Sewel 1 Concentrator with its old Teniente 5 haulage system. The Col6n Concentrator will be fed from the Central and South Mines. Figure 1 shows the transfer systems and ore flows. The Ore Body In plan view at the Teniente 4 level the ore body has a kidney shape 1.800 m long and between 400 to 700 m wide. The north-eastern section of the ore body was washed away by the Teniente River. At the Sub 5 Level, 125 m below Teniente 4, the mineralization completely surrounds the circular 1.000 m diameter steep-sided barren breccia pipe (Braden pipe) . The ore body rock is 73% andesite, 14% diori te, 13% dacite and 6% mineralized breccia. Most of this ore is secondary enrichment sulfides which apart from the dacite is strongly fractured being chal cocite, cove1ine , bornite and

Jan 1, 1981

By E. G. Parker

"Sherritt's zinc pressure leaching process converts zinc sulphide concentrates directly to zinc sulphate solutions and elemental sulphur. Thus, one pressure process step replaces both the roasting and leaching operations of the conventional roast-leach route. In addition, high single-pass extractions eliminate the need for residue treatments to recover unleached zinc. Because pressure leaching produces elemental sulphur, rather than sulphur dioxide, acid plants and smoke stacks are not required, and air pollution control and work-place hygiene are greatly improved. .In 1977, Cominco Ltd. and Sherritt Gordon Mines Limited jointly operated a three-tonne-per-day pilot plant to demonstrate pressure oxidative leaching of Sullivan zinc concentrate and the subsequent separation of elemental sulphur. The process consisted of regrinding the concentrate, continuous leaching at 150°C at elevated oxygen pressures in a four compartment autoclave and the subsequent recovery of elemental sulphur, either in the molten liquid form from a decantation autoclave or as a flotation concentrate. Of the zinc contained in the concentrate, 97% was extracted in 90 minutes, and 96% of the sulphur was converted to the elemental form.Following the successful pilot campaign, Cominco began building the world's first commercial zinc pressure leaching plant. This plant, which was in operation by late 1980, will eventually treat about 25% of Cominco's planned zinc capacity of270,000 tonnes per year.This paper describes the general process principles, the results obtained from the pilot plant and the integration of the process into Cominco's operations at Trail. The application of pressure leaching to a ""grass-roots"" plant is also discussed, and compared with conventional roast-leaching."

Jan 1, 1981

By Joseph A. Gill, Ernest A. Curth

The Government and Leeco concluded a cost-sharing agreement in 1976 to demonstrate longwall mining of a thin coalbed, 1 m or less, in a mine near Hyden, Ky. A premining investigation laid the groundwork for mine design and equipment specifications. Four- legged shields, a face conveyor, a stage loader and an in-web shearer were selected to operate on a 183 m face. Lengwall mining began in 1979 and initially was slowed by low mining height and equipment problems. A rock mechanics program provides early warning capability and criteria for future mine design. Monitoring of surface effects is included.

Jan 1, 1981

By Dell H. Adams

Coal exploration may be defined as the acquisition of data necessary to define and acquire a block of coal which can be mined at a profit. Unlike ore minerals, coal resources are widespread and relatively abundant with supply exceeding demand in almost every case. Thus simple discovery of coal is generally not enough. The chemical and physical variability of individual coal seams, the capital intensive nature of coal mine development, and the changing patterns of utilization require that a multiphase, three-dimensional evaluation be made of each potential reserve. Factors which must be investigated include not only the quantity and quality of the coal but also the geologic, hydrologic, and engineering characteristics of the seam and its surrounding strata. Acquisition of property to provide a secure base for exploration and mine development is equal in importance to the geologic and engineering programs. The successful development of a coal reserve requires that control of coal and, in the case of a strip mine, the surface be in the form of contiguous blocks of sufficient size to support mining.

Jan 1, 1981

This publication summarizes, for all interested parties, the research, development, and demonstration in-house projects programed by the Bureau of Mines for fiscal year 1981 (October 1, 1980 - September 30, 1981) under its Minerals Health and Safety Technology program. The objective of these projects is to provide an ordered and sequenced series of advances toward the Bureau's overall goal of providing the system technology required to create a healthier and safer working environment for the Nation's mining and minerals processing workers.

Jan 1, 1981

By M. J. Sapko

A moderately scaled apparatus was developed by the Bureau of Mines to determine the fire resistance characteristics of mine conveyor belts and similar type materials. The design of the apparatus was based upon data obtained in full-scale fire tests and features a radiant panel to preheat the belt and a methane-oxygen ribbon burner to ignite the sample under ventilating condi¬tions in a rectangular duct. The test method overcomes the limitations of existing laboratory-scale methods and provides a measure of both ignitability and flammability in quantitative terms. Flame propagation depended upon such variables as air velocity, ignitor heat flux, and belt width and height in the test chamber. Data are presented for nine belt materials and fire resistance ratings are proposed in terms of the flame spread rate, heat release rate, and the critical ignitor heat flux. The ratings discriminate between the fire hazard posed by different fire-resistant belts and tend to be in reasonable agreement with those obtained in full-scale fire tests.

Jan 1, 1981

By J. P. Mogan, E. D. Dainty, A. Lawson

"The toxicity reduction of emissions from two diesel engines by the emulsification of water into fuel oil was investigated to determine the applicability of such systems to underground mining machinery. . The engines studied were: 1) the Deut; F6L7J4, four-stroke, air-cooled, 6-cylinder, V-configuration, indirect-injection (IDl) engine, and 2) the Detroit Deisel Allison Division 8V7JN, two-stroke, water-cooled, 8-cylinder, V-configuration, direct-injection (Dl) engine.The application of emulsified fuel treatment to the two-stroke engine did not prove to be practicable. In the case of the Deutz engine, however, substantial benefit resulted. For a modest J5 per cent water content in the fuel and an injection timing of24° BTDC, approximate reductions of40 per cent in the concentrations of both particulate matter and nitric oxide (NO) were produced. These reductions were accompanied by insignificant changes in other toxic constituents. The net result was an approximate 30 per cent reduction in the over-all exhaust toxicity for typical LHD operation as assessed by the Health Effects Index criterion. This toxicity reduction represents a significant advance from the points of view of an improved underground environment and/or reduced ventilation costs."

Jan 1, 1981

By Henry T. Miller

INTRODUCTION The basic purpose of this paper is to suggest procedures and practices to insure that no employee working in a uranium milling operation receives exposure to radioactive, toxic, or other materials or agents that might produce a permanent, deleterious effect upon his physical health and well-being. This program is also designed to insure that each employee can carry out his assigned duties without risk to his health or to that of his fellow workers. The assumptions underlying this program are by necessity, conservative, with emphasis being placed upon prevention rather than correction or reaction to an undesirable situation. The limits specified-either as Threshold Limit Values (TLVs), Maximum Permissible Concentrations (MPCs), or Radiation Protection Guides (RPGs)--represent both standards of best practice and those of law and regulation. The total program is envisioned as a balanced combination of occupational hygiene and radiation monitoring. This includes surveying, air sampling, personnel dosimetry, bioassay, medical surveillance, epidemiology, and training--all backed by a thoroughly tested and evaluated set of emergency procedures. The program, as presented, is keyed to the results of monitoring, surveying, air sampling, medical surveillance, and epidemiology--it being obvious that no problem can result when no hazard can be identified. The lack of an identified hazard, however, does not negate the need for such a program and prudence suggests the need for routine surveillance programs in the best-designed facility. [The Model Mill] The model mill used in this paper is that presented by the U. S. Environmental Protection Agency in its study of the uranium fuel cycle1. The model mill processes 600,000 metric tons of ore per year into 1,140 metric tons of U308 as yellowcake per year using an acid leach process. Other characteristics of the mill such as location, tailings dam design, number of light water reactors supported, and geographical location are not pertinent to this paper. The usual remote western location of these facilities produce some difficulties in delivering the required occupational medical services and is not within the scope of this paper. The operating lifetime of most mills is commonly estimated as from 12 to 15 years in actual practice; however, the U. S. Environmental Protection Agency estimated a 20-year lifetime for its model. The principal steps in the acid leach process include :2 (a) Ore is blended and crushed to pass through a 2.5 cm (one inch) screen. The crushed ore is then wet ground in a rod or ball mill and is transferred as a slurry to leaching tanks. (b) The ore is contacted with sulfuric acid solution and an oxidizing reagent to leach uranium from the ore. The product liquor is pumped to the solvent-extraction circuit while the washed residues (tailings) are sent to the tailings pond or pile. (c) Solvent extraction is used to purify and concentrate the uranium. (d) The uranium is precipitated with ammonia and transferred as a slurry. (e) Thickening and centrifuging are used to separate the uranium concentrate from residual liquids. (f) The concentrate is calcined and pulverized. (g) The concentrate or yellowcake is packaged in 208 liter drums for shipment. This process is illustrated in Figure 1.3

Jan 1, 1981

By J. Ungar, F. X. McGarvey

The selectivity of ion-exchange sites varies with the matrix of the polymer, the nature of the site. and the method of application. The effect of these variables on the recovery of uranium is examined. with particular attention to vinyl pyridine and to symmetrical and asymmetrical quaternary ammonium sites. The effect, on the loading. of high concentrations of chlorides in the liquors is discussed. and other components of the liquors that may cause fouling of the resin are also considered. The discussion is supported by the results of tests conducted on various types of resins. The economics of the process is discussed briefly.

Jan 1, 1981