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By W. H. Patnode

The effect of suspended solids on the resistivity of slurries is discussed and the relationship between drilling mud resistivity and mud filtrate investigated. It is concluded that it is erroneous to substitute mud resistivity for mud filtrate resistivity in electric log calculations. A recommendation is made that both the bud resistivity and the mud filtrate resistivity be determined when electric logs are run. INTRODUCTION The electric log is influenced not only by the resistvity of the drilling mud in the borehole at the time of logging but also by the resistivity of the drilling mud filtrate. Sherborne and Newtoni investigated the relationship of mud resistivity to mud filtrate resistivity and concluded that, "The resistivity of the mud in most cases closely approximates that of its filtrate," and "In fact, with the exception of Aquagel and its filtrate, the figures for any particular mud and filtrate are almost identical." Present practice is to determine only the drilling mud resistivity and apply this same value to calculations involving the mud filtrate. The purpose of this study is to reexamine the factors governing the relationship between mud resistivity and mud filtrate resistivity. EFFECT OF BOREHO1.E FLUID ON THE ELECTRIC LOG Resistivity Log The resistivity log may be modified by the resistivity of the borehole fluid in two different ways: (1) The apparent resistivity of a for-formation may be different from the true resistivity of the formation because of the flow of some current through the drilling mud in the borehole. Therefore the resistivity of the mud is an important factor. (2) The apparent resistivity may differ from the true resistivity, if a formation is invaded by mud filtrate, because of displacement by the mud filtrate of some of the interstitial fluid in the formation. In this case the resistivity of the mud filtrate rather than the resistivity of the mud is the important factor. Self Potential Log The self potential arises, in part, from electrochemical effects resulting from the interaction of connate waters in porous formations and the fluid in the borehole. Expressed in simple form, E = Klog-p where E is the electrochemical self potential, K is a derived constant, pl is the resistivity of the borehole fluid, and p2 the resistivity of the water in the formation. A theory of the electrochemical component of the self potential in boreholes has been recently set forth by Wyllie.3 In the above equation resistivities have been substituted for activities of the ions in the fluids.' It is therefore apparent that the resistivity of the mud filtrate is more nearly representative of the activities of the ions than is the resistivity of the mud. However, it is possible that in some instances the ionic activities of cations from certain clays may contribute to the total cationic activity of the drilling fluid to such an extent that the mud resistivity is more nearly representative of the activities than the filtrate resistivity. This is particularly the case when the resistivity of the mud is less than the resistivity of the mud filtrate. In addition the apparent self potential may be influenced by the resistivity of the drilling mud because of current flow through the borehole. RESISTIVITY OF SLURRIES Aqueous drilling muds are slurries containing fine-grained solid particles. The solid constituents consist mainly of added clays and weighting materials in addition to solids contributed by the drilled formations. The filtrate is primarily water in which quantities of salts or other chemicals are dissolved. The resistivity of the fiiltrate is a function of the type and quantity of dissolved material whereas the resistivity of the mud is a function of the combined resistivities of the filtrate and the resistivities of the suspended solids. Experiments have been carried out to determine the relationship between the resistivity of solutions and the quantity and type of solid matter insus-pension. Solid materials of high resistivity, as well as solid materials of relatively low resistivity, have been used. The data obtained make possible the evaluation of the probable effect of suspended solids on the resistivity of drilling mud. Procedure Resistivities were determined by means of a conventional conductivity cell with platinized-platinum electrodes. Total resistance between the electrodes was measured by Kohlrausch's alternating current bridge method using a General Radio Company Type 650-A impedance bridge with telephone. The cell was standardized with potassium chloride solutions of known normalities in order to calibrate the cell so that measured resistances of slurries could be converted to resistivities. Resistivities were determined for mixtures of potassium chloride solution and solid materials by placing a measured quantity of solution in the cell and adding weighed quantities of solid materials in small increments to the solution. The net change in resistance on addition of solid materials was measured. Even distribution of the solid particles was maintained within the cell by a motor-driven glass propeller before measurements were made. Slurries Containing High-Resistivity Solids Powdered silica sand having a maximum diameter of about 60 microns and precipitated chalk of commercial grade were used to make the slurries whose resistivities were measured. Both of these substances have high resistivities, are virtually insoluble, and effectively do not carry current in a slurry. The resistivities of slurries composed of potassium chloride solution and these two solid materials are given in Table 1. The ratio of the resistivity of the solution to the resistivity of the slurries was computed and was found to follow the relationship established by Archie

Jan 1, 1949

By W. Bonfield

The temperature dependence of the microscopic yield stress (the stress to produce a plastic strain of 2 x 10-6 in. per in.) and the stress-plastic strain curve of polycrystalline Cu 1.9 wt pct Be have been measured for the solution treated condition, an intermediate condition containing G.P. zones and ?' precipitate and the overaged ? precipitate condition, in the range from -58° to 200° C. A transition in micro -yield behavior and a large temperature dependence were noted for the intermediate condition, which are interpreted in terms of the interaction of glide dislocations with two differently sized zones. In comparison the microscopic yield stresses of the solution treated and overaged conditions were less sensitive to temperature variations and are satisfied by the Mott-Nabarro and dislocation bowing theories, respectively. A determination of the temperature dependence of the yield stress of a precipitation hardening alloy has provided a powerful tool for evaluation of the operative deformation mechanism. There is a marked contrast between the effect of temperature on the yield behavior of a metal containing coherent zones or intermediate precipitates, which can be "cut through" by mobile dislocations, and a metal containing a dispersion of noncoherent particles, through which dislocation "bowing out" is the dominant role of deformation.' These studies have in general been confined to single crystals, as it was considered that similar experiments on polycrystalline material did not produce good data because of the lack of sensitivity with which the yield stress could be determined. However, this objection has been removed by the introduction of mi-crostrain techniques, with which the yield stress in polycrystalline materials can be measured to a strain sensitivity of 10-6. Such measurements have not only shown that the deformation of polycrystalline precipitation hardening alloys can be examined with the same detail as single crystals, but also that some unexpected results are obtained.' In this paper the results obtained from a study of the temperature dependence of the microscopic yield stress (the stress to produce a plastic strain of 2 x 10-6 in. per in.) and the stress-plastic strain curve of a polycrystalline Cu 1.9 wt pct Be precipitation hardening alloy (Berylco 25) are discussed. The temperature dependence of the alloy was measured for three different conditions: 1) The solution treated condition (a supersaturated solid solution of a containing ~12 at. pct Be3) which is obtained by water quenching the alloy from 800° C. 2) The condition of y' intermediate precipitate, to- gether with some G.P. zones,' which is produced after an aging treatment of 2 hr at 315°C from the solution treated condition. (The alloy was cold rolled to 40 pct reduction prior to aging to minimize grain boundary precipitation effects.)4 3) The condition with equilibrium ? precipitate structure2 which is developed after an aging treatment of 24 hr at 425° C. EXPERIMENTAL PROCEDURE Tensile specimens of gage length 1 in. and with rectangular cross section of 0.18 by 0.06 in. were prepared from the solution treated, cold rolled alloy and were either resolution treated for 1 hr at 800°C, followed by water quenching, or aged for 2 hr at 315°C and 24 hr at 425° C to produce the desired precipitate structures. The microstrain characteristics of the aged specimens were determined at temperatures from —58" to 200° C and those of the solution treated specimens from -58° to 30° C. Each temperature was controlled to ± 0.2°C, which was a level of stability sufficient to eliminate thermal expansion effects from the measurements (~1.2°C temperature increase produced an extension of 2 x 10-6 in.). The microplastic behavior of the specimens in the temperature range below 82" C was measured with a standard Tuckerman strain gage,5 while at temperatures above 82°C a modified Tuckerman gage with a reduced strain sensitivity (4 x10-6 in. per- in.) was used. A load-unload technique was used to establish values of the microscopic yield stress. The specimen was strained at a constant cross head speed of 2 x 10-2 in. per min to a given stress level, at which the total strain was measured. Then the specimen was immediately unloaded at the same rate and any residual plastic strain determined. This procedure was repeated for an increasing series of stress levels until the microscopic yield stress was established by a direct measure of the stress to produce a residual plastic strain of 2 x 10-6 in. per in. (It should be noted that, as reversible dislocation motion occurs at stresses less than the microscopic yield stress,2 the plastic strain rate at this level was not constant.) In an ideal test, the microscopic yield stress would be determined from a continuous stress-strain measurement, rather than from a load-unload sequence, in order to eliminate mechanical recovery effects.6 However, it was found experimentally that mechanical recovery was negligible in Cu 1.9 wt pct Be at small plastic strains for all the temperatures investigated, as the microscopic yield stress was independent of the number of load-unload cycles employed (i.e., the values measured for specimens subjected to different numbers of cycles was within the experimental scatter determined for specimens tested in an identical manner). Therefore, it is reasonable to consider the microscopic yield stress determined in the load-unload

Jan 1, 1969

By James R. Stuart

AS the available remaining properties of iron ore reserves on the Mesabi Range are opened up for mining, the various properties located under lake beds are brought nearer an active status. The actual physical problems involved in stripping these properties do not act as a deterrent so much as the legal and political problems that are encountered. When it is proposed to destroy a natural lake that has been used by the public for many years, much local as well as state opposition may be encountered regarding its destruction. Public hearings must be held and some adverse publicity is likely to result. The ownership of the ore under the lake and the rights of the abutting property owners must be settled, and protection from damage caused by a disturbance in surface and subsurface drainage is likely to be demanded by property owners some distance from the proposed mine area. The Embarrass Mine, located near Biwabik, Minn., falls into this classification. A portion of the orebody lies under what was formerly Syracuse Lake, this body of water having been removed in the process of stripping the mine. An additional problem in the case of a meandered body of water is the establishment of a meander line that can be projected downward as mining progresses to form the basis for a satisfactory division between lake bed and upland ore shipments for royalty purposes. Fig. 1 illustrates the complications encountered in maintaining these divisions. A balance point was agreed upon in the center of the lake to make an equable division of lake bed ore to the abutting properties. The entire lake bed has since been adjudged the property of Minnesota. Lake Characteristics Lake bed stripping problems with which this paper is concerned necessarily are limited to a specific type of lake, namely the glacial lakes of the Lake Superior region. One characteristic common to these bodies of water is a deposit of fine black mud or silt on the bottom, frequently underlain by a layer of impervious blue clay. This is also true of the muskeg areas of the region, which present almost identical problems as lakes in stripping. The actual removal of the water and the lake bed material is a routine matter more or less standardized as to equipment, and the period of time required can be estimated easily on the basis of volume and capacity. More important than the foregoing is the execution of preliminary work, and above all, the timing involved. An account could be prepared based entirely on statistical and cost data which would give a very fair picture of the time required and cash outlay needed to effect the removal of a body of water preliminary to stripping the orebody. However, the real interest from the standpoint of the operator and the engineer who carry responsibility for completion of the job lies in the unexpected emergencies and the action of various materials involved in the stripping when the balance has been upset through diversion of water courses and the reduction of the lake level. Runoff and Drainage Lakes are located in natural basins that catch all the rain water and runoff water for a considerable area. Where a lake is involved having an inlet and outlet or a sizeable water course running through it, the drainage area may include a watershed covering many square miles. All available data then must be collected to supply a history extending over as many years for which information can be gathered on the flow of streams, annual rain and snowfall, and most important, the peak flows to be expected. Where the diversion of a stream around the stripping area is a part of the problem, this last factor is of great importance since it controls the cross-section to be selected for the diversion channel and the volume to be removed in its excavation, as well as affecting the hydraulic considerations to be met in the design of the completed channel. Characteristic material in the overburden found at the Embarrass Mine is illustrated in Fig. 2. Well Pumping Pumping from the well holes was started well in advance of the draining of the lake. Fig. 3 shows a gradual lowering of the water table with no noticeable fluctuations during the period in which the lake was being dewatered. Unfortunately, because of tight ground, a maximum flow to the wells was not maintained. This retarded the rate at which the water table was reduced so that in the course of stripping the excavation soon extended below the water table, and the great bulk of the pumping was handled from a system of sumps in the pit itself. Any dewatering program projected by prepumping from wells, a glorified well point system, would have to be started well in advance of the stripping to be of any great advantage. Preliminary drainage of the surface over the mine area is entirely apart from the actual elimination of the lake bed itself. Since the lake is what is called a perched water table because of the impervious character of the lake bottom, the adjoining surface may be dewatered below the surface of the existing lake and the flow will not be affected by the proximity of that body of water. This condition actually has been demonstrated through the establishment of a number of observation holes where a small churn drill was used to put down the holes and a 3-in. pipe was installed for taking water level

Jan 1, 1952

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Jan 1, 1938

By Javier Ramirez

"Recent R&D pilot plant testing was done to validate a full industrial scale system for recovering residual magnetite from the tails of an iron ore concentration plant. The recovery system is based on an autogenous heavy media separator treating a tailings slurry at 0.45% solids by weight. A 3,100 m3/h capacity concentrator was installed based on the testing findings. Excellent metallurgical results have been observed in two years of continuous operation. This experience is the basis for the proposal of a novel process consisting of the preconcentration of base metals ores; the idea being to obtain a heavy bulk concentrate of the polymetallic minerals and discarding in one step the light gangue mineral components. The bulk concentrate then proceeds to conventional froth flotation. The reduced volumetric flow of the bulk concentrate allows a reduction in the size of the subsequent flotation circuit. A prerequisite for the heavy media separation is a fine size grinding of the ore and a correspondingly high degree of liberation of the mineral species to be recovered. A process scheme is also presented in which the gravity concentrator works in conjunction with a companion tailings thickener for process water recovery. The underflow from this thickener, having paste consistency, can be sent to a tailings dam or mixed with Portland cement for mine backfill. The process water in the system is managed at an estimated recovery in excess of 98%. Additional R&D work using this novel approach is being considered by a large iron ore concentrator for removal of liberated silica from the final concentrate; and for upgrading 3.5 million tons of a marginal hematite reserve owned by the mining company.INTRODUCTIONGravity preconcentration (static and dynamic) of sulphides prior to froth flotation has been and is being practiced at various concentrator plants (Asarco´s Young Mill: heavy media cone; Mount Isa: DSM cyclone; Afton Mines: Denver jigs; Tsumeb Mine: Reichert cone) (Burt, 1984). The idea behind installing preconcentration equipment is to reduce the volume of feed processed on flotation, thus requiring smaller flotation plants although other advantages may be of importance, such as obtaining richer concentrates due to higher grade feed, better process water and energy management, and reduced flotation reagent consumption.The SedconTM solid/liquid separator is a versatile process unit patented in Mexico (MX/u/2007/000305). As a gravity device, it may be classified as an autogenous heavy media dynamic concentrator. Figure 1 depicts the Sedcon (Sedimentador de conos) separator. The four main components are as follows: the feedwell for efficient mixing of the pulp with dilution water and flocculant (and coagulant for water clarification); the inner cones (passive elements for solid/liquid separation); the discharge underflow cone; and the scraper mechanism. The standard discharge cone is a 60° inclined cone for general applications. For paste tailings thickening applications, a deep cone with a 69° inclined cone is used."

Jan 1, 2016

By Niels Hansen

The substructure formed by drawing at room temperature in aluminum (99.5 and 99.998 pct purity) and in recrystallized aluminum aluminum-oxide products containing from 0.2 to 4.7 wt pct of aluminum -oxide was examined by transmission electron microscopy, and the flow stress of the drawn materials was measured by tensile testing at room temperature. A sub-grain structure was present after a reduction in area by drawing of 10 to 20 pct, and the subgrain size was observed to decrease with increasing deformation. The tensile data show that the increase in flow stress (0.2 pct offset) by drawing from 10 to 95 pct depends on the reduction in area, not on the composition of the materials. Dispersion strengthening and subgrain bowzdary strengtlzening contribute to the flow stress, and these strengthening processes have been found to be linearly additive. The flow stress (0) can be related to the subgrain size &) by the Petclz relation = uo + k . t, where go is dependent on the composition of the products and k is approximately the same far all materials. THE microstructure of dispersion-strengthened aluminum aluminum-oxide products consists of small oxide plates distributed in an aluminum matrix. The matrix structure depends on the manufacturing history, and in hot-worked as well as cold-worked products the matrix is divided by subgrain (or dislocation) boundaries. For hot-extruded products it has been shown1 that dispersion strengthening and subgrain boundary strengthening are linearly additive, and the flow stress (0.2 pct offset.) at room temperature has been related to the subgrain size (ts) by a Petch equation,2,3 s = so + k . ts-1/2, where s0, increases with increasing oxide content. For cold-worked products containing subgrains no systematic work has been reported, and it was the aim of the present study to examine the microstructure and the relationship between the flow stress and the subgrain size for such products. The behavior of' aluminum aluminum-oxide products depends on the purity of the aluminum matrix, and aluminum of the matrix purity (99.5 pct) was included in the investigation. The literature contains few data about the behavior of this impure aluminum, and aluminum of a higher purity (99.998 pct) was therefore also examined. As regards the relationship between the flow stress and the subgrain size in cold-worked dispersion-strengthened products, no systematic work has been reported. For aged cold-worked structures containing fine precipitates (Fe-Mo carbide) a Petch relation has been found,4 and it has been shown that the k value NlELS HANSEN is Head, Metallurgy Department, Danish Atomic Energy Commission, Research Establishment Riso, Denmark. Manuscriot submitted January 9, 1969. IMD is approximately the same as in iron, whereas the s0 value is higher owing to the presence of the precipi-tates. Investigations of metals such as tungsten,' ferrous metals,4,6 and molybdenum7 cold drawn or swaged at room temperature have shown that the flow stress can be related to the subgrain size by a Petch relation when ts is taken as the subgrain size perpendicular to the direction of deformation. For aluminum no work has been reported on the relationship between the flow stress and the subgrain size after deformation at room temperature, whereas for aluminum tensile strained at different temperatures in the range -183" to 375°C a Petch relation has been found by taking ts equal to the subgrain size.' In the present study two aluminum materials (99.998 and 99.5 pct) and three aluminum aluminum-oxide products (containing 0.2, 1.0, and 4.7 wt pct oxide) were drawn at room temperature to reductions in area from about 10 to about 95 pct. The structures were studied by transmission electron microscopy, and the flow stress (0.2 pct offset) was measured at room temperature. EXPERIMENTAL Materials. The materials are given in Table I together with the chemical analysis. The three aluminum aluminum-oxide products were manufactured from aluminum powder that had been compacted and Table I. Chemical Analysis of Materials Al203 Fe SI Material wt pct wt pct wt pct 99.998 pct* - 0.0004 0.0012 Aluminum 99.5 pctt - 0.36 0.16 Aluminum AlMD13† 0.2 0.16 0.12 Aluminum- -Oxide AlMD105† 1.0 0.26 0.18 Products SAPISML960s† 4.7 0.22 0.19 *Other impurities: O.0004 pct max each of Cu and Zn (supplier's analysis). †Other impurities: 0.03 pct max Cu. 0.02 pct max each of Mn, Mg, Zn, Ti. Table 11. Mean Diameter of Aluminum-Oxide Particles in Extruded and in Cold--Drawn Aluminum Aluminum-Oxide Products Mean Diam. of A1203-Plates* Material State A AlMD105 Extruded 540 Cold Drawn 97 pct 510 SAP ISML 960 Extruded 770 Cold Drawn 95 pct 820 *The standard deviation of the mean is approx. 5 ±pct.

Jan 1, 1970

By Rudolph G. Wuerker

MUCH descriptive matter has appeared on the subject of suspension roof supports, or roof bolting, as it is more commonly called. The widespread introduction of roof bolting into coal mines and metal mines is truly phenomenal. Mine operators were quick to recognize the advantages of supporting wide openings without hindrance to machine maneuverability and ventilation. Although suspension roof support has long been installed at St. Joseph Lead Co. mines in southeast Missouri,'" its application to coal mining presented new problems, such as proper anchorage and bearing for the bolts, bolt diameter, and spacing of bolts. After continuous testing and experimenting at the mines, standard roof-bolting materials were determined.'!' The study reported in this paper is not concerned with such details as bolt diameter, which may be considered already solved in practice. In the tests discussed here, small models patterned on actual bolts were found to function in the same way and as satisfactorily as their prototypes. The aim of these tests was rather to investigate the influence of roof-bolting systems on the stress distribution around mine openings and to study the fracture patterns obtained in actual testing. Little was found about this in the literature, as testing of suspension roof methods and quantitative measurements are only now coming to the fore. Several suggestions and actual measurements have been made to evaluate critically the functioning of roof bolting systems, single roof bolts, and parts thereof. Outstanding among them is Bucky's outline of structural model tests.'" Since none of the suggested testing equipment was available, however, for the experiments discussed below, a different approach was chosen. The response of a mine roof under stress has often been compared to that of a beam. The slow coming down and bending through of beam or plate-like banks of shale, sandstone, or top coal is a familiar occurrence, extensively cited in the literature." It was felt that testing of roof-bolt systems installed in a concrete beam which was loaded in bending would be a fair approximation of the behavior of a mine roof underground. Another school of thought considers the roof behavior over an underground opening in connection with the stress distribution all around a circular or rectangular opening. This is more accurate, and leads to the concept of a dome-shaped zone of material destroyed under tensile stress. This is likewise a common sight in unsupported roadways where the continuous fall of roof results in what has been called the natural outline of roof fracture. This theory could not be tested and is treated separately in Appendix B. It is important to note that according to both assumptions the immediate roof fails in tension; the use of a beam in these tests, therefore, should give information valid for either of the two theories. With the testing equipment at hand it was possible to load concrete beams 6xlx0.5 ft under two-point loading, giving an equal bending moment over the center part in which the model bolts were installed. A comparison was made of the ultimate loads needed to break plain beams and beams in which roof bolts were installed. Arrangements were made with: 1—plain beams; 2—bolts with plate washers, some with holes drilled at 90" angles and others with holes drilled at 45" angles; 3—bolts with channel irons underneath; 4—bolts in holes filled afterward with cement; and 5—bolts anchored in a stronger stratum. The foregoing arrangement is made in order of increasing strength, as assumed from the theory of reinforced concrete. Likewise, laminated beams with wooden model bolts and with combinations of the foregoing set-ups were tested. All in all, 21 experiments were made out of the much greater number of combinations possible. There were, too, some trial tests. Enough observations from this limited number were made to interpret the behavior of mine roof, supported by various types of suspension bolts, at fracture. In present-day concepts, which have been proved by mathematical derivations and stress analyses, any opening driven underground will change the distribution and magnitude of the stresses existing around it. It does not matter whether the stresses become visible, as in rocks whose strength is less than the forces acting upon them, or whether they are invisible, as in the gangways lacking evidence of rock pressure. In this latter case the rocks can withstand changes in stress-distribution. To consider the mine roof as a beam, there are, with transversal loading, tensile stresses in the lower fiber and compressive stresses in the upper layers above the neutral axis of the beam. Beams of brittle material such as rock and concrete fail exactly as shown in Fig. 1. Nearly all model beams showed the same fracture pattern as that of a tension crack. The influence of support, by roof bolting or conventional

Jan 1, 1954

By Glenn C. Waterman

THE diamond drill is a very important tool in exploration and development testing and its use is increasing. In almost all cases results of diamond drilling are analyzed on the basis of grade and tons. A proper evaluation of core and sludge assays is important if drilling results are to be acceptable as a basis for geologic and engineering appraisal. The relatively wide variation in assay averages as calculated by various well-known combining methods indicates that the engineering choice of a method may affect the outcome of the drilling in terms of ore and waste. The problem of combining assay results from core and sludge samples has been discussed many times in conference and in the literature.'-' Most writers agree that the field of disagreement in methods is large and that the engineer on the job must consider features unique to his drilling, pick one of several combining methods, and depart from the rules when abnormal results come in. All the discussion to date can be summed up by the admission that as yet there is no fairly simple, generally acceptable combining method that is practicable over a wide range of drilling conditions, ground conditions, and ore occurrence. The combining problem is important in evaluating drilling results at Chuquicamata. Recently a reappraisal has been made of recovery variables and their effect on assays, with the result that a new combining method is offered which fits average drilling conditions and is mathematically reasonable. It is simple in application, fundamentally correct, and an improvement over most combining methods. At Chuquicamata diamond drillholes are used to outline the grade and position of blocks of normal and marginal grade oxide, mixed, and sulphide ore. Most holes penetrate all classes of material (and waste), and it is important for mining programs as well as ore reserves to know almost precisely the soluble and insoluble copper content of mineralized ground. At present three classes of ore are mined and treated differently. For an orderly sequence of mining operations which can provide regular daily tonnages of all three ore types and keep grade at certain levels with minimum variation, ore type and its grade must be predicted. Diamond drilling plus geologic mapping and bench sampling are tools for prediction. And drilling data are largely used to calculate grade of material more than a few meters away from bench faces. The orebody at Chuquicamata" is criss-crossed by millions of barren or mineralized weak to fairly strong slips and fault fissures. Mineralization is diverse and encompasses many quartz and oxide or sulphide-bearing copper veins, as well as seams and disseminated grains. Copper occurs in oxide or sulphide minerals or mixtures of these two mineral types. Rock conditions vary from intensely seri-citized (soft and porous) through clay-altered ground to almost fresh granodiorite. The result is an orebody which offers many obstacles to good and consistent core recovery in diamond drilling. Recovery varies considerably in the several alteration zones, the various types of oxide and sulphide ores, and the position and inclination of the drillhole within the complex fracture pattern. As core recovery drops sludge samples must be used with core samples to calculate grade. Many years of drilling at Chuquicamata indicate that in good grade oxide zones and in the sulphide areas core recovery is good and the ground uniformly mineralized. Moderate loss of core, therefore, does not markedly affect grade calculations based on core assays. An early core-sludge combining method used core assays at face value as indicating grade down to 50 pct core recovery, but below this recovery percentage sludge samples were used and weighted according to the standard Longyear chart. This method apparently did not introduce serious errors, but it abruptly used sludge assays with high weighting factors representing 100 pct return irrespective of actual percentage of sludge recovered. Recent drilling activities have been directed toward outlining the marginal ore areas. The non-uniform mineralization and generally poorer core recovery in such ground indicated that a more exact core-sludge combining method was required to equate wide differences between core and sludge assays and recovery. In fringe ore areas at Chuquicamata core recovery averages about 50 pct, from a minimum of 10 to 15 pct to a maximum of 100 pct. Sludge recovery is likewise variable and averages perhaps 80 pct even though holes are cemented as water return falls off. In a homogeneously mineralized area cut by many slips and faults, with hard and soft ribs, loss of core is loss of ground which has a grade similar to that of core recovered, and core assays approximate true grade. In this case sludge samples need not be used. However, it would be unusual to know beforehand that an area is uniformly mineralized, and in fact this condition is probably uncommon. Generally the distribution of valuable minerals in the ground does not exactly compare with their recoverability in core. Thus, in the usual case, loss of core decreases the

Jan 1, 1956

By O. D. Gaither

This paper is an analytical study of the flow of fluids through small vertical conduits. Small conduits are defined as 11/4-in. nominal diameter tubing size and smaller, and approximately twice this area for annular conduits (i.e., 1- X 21/2-in. annulus and smaller). Experimental data are presented for the 1-X2-in. and 11/4- X 2%-in. annuli, and the I-in. and 11/4-in. tubing, since these represent the small conduit sizes and configurations generally encountered in oilfield applications. Data have been gathered for these conduits for single-phase water, single-phase gas and two-phase water-gas mixtures, with particular emphasis on high gas-liquid ratios. Water rates in excess of 2,000 BID and gas rates in excess of 2.5 MMcf/D, and two-phase flow ratios in between these two, represent the scope of the data gathered. Existing equations have been applied to predict flowing pressures and compared with experimental data. New correlations have been developed. INTRODUCTION The increased economic pressure on the domestic oil industry in the United States has constantly required the use of new techniques and equipment designed to reduce the cost of finding and producing oil and gas. Since tangible items are most readily apparent in economic analysis, the advent of lower-cost well completions was inevitable. One of the methods used to reduce costs which has received widespread attention is the slim-hole completion technique where tubing is used as the well casing and in which small conduits are used for tubing if necessary. Small conduits, defined by Kirkpatrick1 as "11/4-in. diameter nominal tubing and smaller for tubing flow and less than twice the 11/4-in. diameter nominal tubing internal flow area for annulus flow", have also found widespread usage as siphon strings for de-watering gas wells and as "kill" strings in deep high-pressure oil and gas wells. The growing use of small-diameter tubing has resulted in an increased need for development of improved methods to measure or predict flowing bottom-hole pressures since the physical dimensions generally preclude the use of subsurface-recording pressure gauges. Even in the cases where small bombs are available, the relatively high velocities encountered at nominal flow rates make it necessary to use excessive weight bars or special hold-down devices. Attempts to use recognized correlations to accurately predict flowing or gas-lift performance in wells equipped with small conduits have been generally unsuccessful. Insufficient field data were available to allow the development of a correlation on this basis, and an experimental approach was applied in an attempt to obtain a workable relation. The experimental approach used to obtain the data presented in this paper was actually a compromise between a field installation and a laboratory study. A test well 1,000 ft in length was used to obtain flow data on single-phase liquid, single-phase gas and two-phase water-gas flowing mixtures. Liquid rates up to 2,200 B/D and gas rates up to 3 MMcf/D were used in the single-phase flow studies. Two-phase flow rates from 100 to 600 B/D with gas-liquid ratios from 500 to 8,000 cu ft/bbl were recorded. Experimental data were obtained for single- and two-phase flow through 1-in and 11/4-in. nominal tubing, and through the annuli between 1- and 2-in. and 11/4- and 2%-in. nominal tubing strings. Experimental results for the two-phase flow are compared to the Poettmann-Carpenter correlation2 which is widely used as a comparative standard for development of multiphase flow predictions in flowing and gas-lift wells. Correlations developed by Tek,3 Baxendell and Thomas" were also investigated. The experimental data recorded herein fell in between the two flow regimes as defined by Ros," and this correlation also failed to yield satisfactory results. The fact that existing correlations failed to confirm the experimental data led to the need for development of a new correlation. Although a two-phase flow study was the primary objective of this investigation, data were also recorded for single-phase flow of water and gas, and constants were developed relating to pipe roughness and equivalent diameters for annular flow. These single-phase studies assisted materially in the development of certain of the two-phase flow results. Considerable previous work has been published which presented relationship of surface measurements to bottom-hole condition. The works of Buthod and Whiteley,6 Jones,' Poettmannb and the Texas Railroad Commission" are classic examples of the successful use of mathematical relationships which allow acceptable predictions of subsurface pressures, when gas is the flowing fluid. Darcy and others have derived relationships which may be used with minor modifications to predict subsurface flowing conditions in injection and water-supply wells. As previously stated, the application of the single-phase flow relationships

By J. H. Viljoen, C. H. Graan, N. B. Strydom, J. S. Greyson

M J Martinson (Fellow) There is everywhere a lamentable dearth of quantitative information on occupational factors affecting the health and safety of mineworkers, and as the authors point out in their introduction underground lighting is one of the factors which has received little attention in South Africa prior to this welcome paper. Since the authors are to some extent pioneering, it is perhaps inevitable that the paper should raise more questions than it answers, but nevertheless it provides a provisional yardstick against which the need for, and success of, remedial action may be judged. Before commenting in detail on a few aspects of the paper it may be worthwhile to summarize the statutory provisions relating to underground lighting, contained in Chapter 15 of the Regulations promulgated in terms of the Mines and Works Act 27 of 1956. Chapter 15 is entitled `Lighting, Safety Lamps and Contraband', but in the context of the present paper only four simple regulations need be noted: firstly, no person shall work or travel in an unilluminated part of a mine unless he carries a light ; secondly, adequate stationary lights shall be provided at stations and certain other localities in shafts, winzes and places used for winding; thirdly, mechanically propelled conveyances shall be provided with an effective bright light shining in the direction of travel; and fourthly, moving machinery to which personnel have access shall be illuminated so as to be clearly visible. In all four regulations the standard of lighting to be provided is wholly subjective, and under these circumstances the wide inter- and intra-mine variations in light intensity disclosed by the study are hardly unexpected. Nevertheless the data presented in the authors' Table I raise three points (1) Is there any evidence that the 19 mines visited during the survey are representative of the 45-50 large gold mines in South Africa, or that the shafts selected are representative of the respective mines? Mine managers would scarcely be human were they not to offer better lit working places for the survey; perhaps the use of a scaling factor (e.g. tonnes or transit mineworkers per shift, kilowatts, etc.) for each working place might have helped to put the figures in perspective. (2) Looking at two of the more numerous - and probably important - of the ten working places included in the survey, i.e. stations and tips, one is struck by the marked. skewness of both intensity distributions (with higher frequencies at the lower intensities) and the abrupt cut-off in both cases just short of pitch darkness. Thus, of the 45 meaned readings recorded at 26 tips no less than 19 fall in the lowest range of 0,1-0,9 1m/ft2, with but one tip in total darkness; of 46 readings on 26 stations, 7 fall in the lowest range and none is in total dark-'1 ness. Although a light with an intensity in the range 0,1-0,9 1m/ft2 may satisfy the regulations, it is certainly arguable that in some situations it may be safer to have no light at all. (3) In view of the skewness referred to above, one wonders how meaningful are the `total average light intensities for (the gold mining) industry' quoted in Table I, which the authors compare with illumina¬tion levels found in other countries. Looking at tips again, one notes that 34 out of 45 meaned intensity readings fall below the `total average' for tips, and only 10 exceed it. From this it may be concluded that a few well-lit working places included in a `total average' may conceal a plethora in Stygian gloom. In their discussion the authors make fleeting reference to the effect of underground lighting on productivity and accident rates, without however quoting any South African experience in this regard. Intuitively it seems highly likely that better lighting (in the widest sense) would improve productivity and reduce accidents, and of course it is in these areas that the mining industry might seek economic justification for providing better lighting. It is all the more regrettable therefore that these factors were not studied simultaneously with the present investigation, especially at the Psychological Division of the Human Sciences Laboratory has recently published a study on underground accidents in the gold mining industry without making any reference whatsoever to lighting as a causal factor. In the event neither of the two studies enables us to make an economic appreciation of the situation, and this seems to be a classic example of a situation where some dovetailing of research effort would have greatly increased the useful return. On the effect of lighting on accidents, it may be noted that in 1939 Mickel presented some very interesting statistics to the Mine Managers Association of the Transvaal2, on the improvement in accident rates which followed the random issue of electric cap lamps to native mineworkers in the Hercules Section of E.R.P.M. the previous year. Based on a 12 month trial period in which 250 000 shifts were worked with carbide lamps and (contemporaneously) 307 000 shifts with cap lamps, the incidence of major accidents (at least one day in hospital among men using carbide lamps was 1,15 accidents per 1000 shifts, while the incidence of minor accidents (requiring dressing only) was 11,52 per 1 000 shifts. The corresponding figures for men using cap lamps were 0,79 and 7,31 accidents per 1 000 shifts respectively, representing a reduction of over 30 per cent in both major and minor accidents. While these figures are obviously dated - only men using cap lamps were issued with hard hats, for instance - they do indicate a need for similar research today.

Jan 1, 1971

By R. R. Klimpel

Introduction The grinding of coal-water slurries has received increasing industrial attention during the last decade. In particular, there is long-term interest in the use of pulverized coal-water slurries to replace oil in combustion equipment and in the development of coal gasification/liquefaction processes that require coal-water slurries as feed. More specifically, the use of coal-water slurries in gasification requires grinding a high-density slurry containing the smallest amount of water consistent with slurry pumping and spraying. As part of a fundamental engineering research support program aimed at the industrial implementation of dense coal-water slurry grinding, this author has published several papers on how specific rates of breakage vary as a function of slurry rheology (Klimpel, 1982,1982/83). These papers demonstrated that there is a consistent pattern of change in specific rates of breakage of coal in dense slurries with controlled variation in slurry rheology. By matching rheological data with laboratory grinding results, it was possible to identify directly slurry conditions that correspond to: 1) slowing down of breakage rates, 2) the occasional acceleration of breakage of some sizes, and 3) conditions where chemical additives will increase rates of breakage. In brief, these conditions were analyzed using two different criteria: a) the net production rate of material less than some specified size (e.g. kg/min of minus 325 mesh) in a standard batch laboratory mill test as a function of controlled changes in grinding conditions, and b) the use of the one-size-fraction feed method, which consists of following the disappearance of this largest size over grinding time in a batch laboratory mill to arrive at well-known specific rates of breakage (Austin et al., 1984). Detailed references to the methodology used as well as the conclusions are available (Klimpel, 1982, 1982/83) and will not be repeated here. The purpose of this paper is to further demonstrate several additional characteristics of dense coal-water slurry grinding that were shown in a simplified sense in the earlier publications of the author but which have clearly demonstrated themselves as being very important in the industrial simulation and scale-up of such coal-water grinding systems. In particular, this includes the clear and unambiguous demonstration of how the simultaneous acceleration of breakage of some size fractions and slowing down of the breakage of other size fractions is occurring as a function of changes in coal-water slurry density. In the earlier publications (e.g. Klimpel, 1982), it was shown by specially designed experiments that the addition of fine material and/or the use of a chemical thickening agent accelerated the specific rates of breakage of coals of coarser size fractions using the one-size-fraction method. There were also numerous examples given of non first-order breakage (the slowing down of coal breakage rates) using also the one-size-fraction method due to the presence of excessive amounts of fines which corresponded to the development of a rheological yield value. The problem with the simulation and scale-up of any laboratory and/or pilot-scale mill data to an industrial scale using the mechanistic modelling approach involving specific rates of breakage and breakage distribution parameters (e.g. Austin et al., 1984) is the number of assumptions involved in translating the smaller mill breakage parameters to the predicted larger mill breakage parameters. It is apparent, at least to this author, that to accurately simulate and predict larger scale equipment performance from smaller scale data (given that the larger scale data performance is known and hence predictions can be thoroughly checked) requires a better knowledge of breakage parameters than is currently available. More specifically, it was felt that one of the chief problems was the inability of the one-size-fraction method of determining breakage parameters to sufficiently represent the actual magnitude and sometimes even the directions of. the complicated interactions involved with slurry density changes in coal-water slurry grinding. Thus, a special set of experiments was conducted in a somewhat larger batch ball mill (0.457 m diam x 0.610 m length) than the 0.203-m-diam mill used in the original rheology characterization paper (Klimpel, 1982) so as to minimize any unusual effects due to wall-ball interactions (2.54-cm-diam balls used in both mills). More importantly, the measurements of specific rates of breakage were done using a proprietary tracer method on a portion of a given size fraction, which was then remixed into a natural feed size distribution before grinding. The experimental procedure and analysis of subsequent data was done in exactly the same manner as the radioactive tracer technique on coals as originally developed by Gardner (1962). The advantage of such an approach is that it makes no assumptions such as the independence of the specific rate of breakage of any size on the absolute sizes and amounts of other sizes present (both larger and smaller) in the mix of natural feed material. It will be shown that the measured rates of breakage using the direct tracer technique and the one-size fraction method on the same coal are indeed different. In fact, an accurate assessment of what is happening to the rates of breakage as a function of changing slurry density can only he made by measuring particle breakage under grinding conditions approximating the size distributions actually being produced in practice. Experiment procedures and results The pilot mill used was 44 cm diam x 60 cm long with a volume of 91,250 cm3 and was fitted with six 0.5-in. lifter bars.

Jan 1, 1992

By C. J. Monahan

INTRODUCTION With the introduction of large diameter hole drilling to underground mining, a need for the development and utilization of new blasting technology has arisen. This new technology has resulted in more efficient blasting and in easier and cheaper mining. The application of spherical charges to the crater blasting method is an im¬portant part of this technology. Spherical charges, or their geometric equivalent, make for highly efficient use of explosives in the cratering application. Their geo¬metrical configuration, normally a length :diameter ra¬tio of <6:1, limits their charge weight size. For exam¬ple in a 165-mm (61/2-in.) hole a spherical charge weighs approximately 34 kg (75 lb). These dimensional and weight restrictions require careful engineering design and control in the production application. Strathcona mine, a large tonnage mechanized cut¬and-fill and blasthole stoping operation located on the north rim of the Sudbury Basin, has successfully em¬ployed the crater blasting method in production blasting for the past two and a half years (as of 1979). As a result, this method, using spherical charges, has made possible the development and implementation of a verti¬cal retreat blasthole mining method which is being used for rib pillar recovery. The crater method used at Strathcona essentially in¬volves blasting off horizontal slices of ore into an under¬cut while retreating vertically to the pillar overcut elevation. Loading and priming of the holes is achieved from the overcut and the blasted muck is recovered by load¬haul-dump (LHD) units through a drawpoint system. Valuable experience on the behavior and performance of explosives in large diameter holes has been gained in both the operating and engineering fields from the appli¬cation of this blasting method. It has been used in drop raising and is currently being evaluated for use in pri¬mary stoping operations. Also, serious consideration is being given to its use for the recovery of post pillars in mined-out and backfilled cut and fill stopes. The method has significant advantages over conven¬tional methods of pillar recovery from economic, ground control, and safety points of view. These include cost benefits due to a minimum stoping development, less ground support, reduced labor requirements, and faster mining rate. Research on the blasting mechanism indi¬cated the method should result in less damage to fill walls than would be normally expected with conven¬tional longhole or large diameter hole cylindrical blast¬ing (benching). This is primarily due to the limited charge weight per delay because of the size dimensions of a spherical as compared with a cylindrical charge and also because the large diameter holes experience less de¬viation than standard-sized longholes. Experience has shown that wall damage has been minimized with the use of this method. The safety aspects of the working environment are also improved as there is no require¬ment for manpower to go under exposed backs after blasting, which is a significant improvement over the cut-and-fill method. CRATERING THEORY General The term cratering, in blasting, is applied to the for¬mation of a surface cavity in a material as the result of detonating an explosive charge in that material. This blasting concept was initially used as a tool in the eval¬uation of explosives performance; however, it has been utilized more recently on surface and to a lesser extent in underground blasting operations. Explosive charges used in crater method are nor¬mally spherical or the geometric equivalent, as research into the application of this breakage mechanism to rock indicates that spherical charges or their equivalent pro¬duce optimum results. In blasting practice, spherical charges have been defined as having a length to diameter (L: D) ratio of 4:1 or less, and up to, but not exceeding a L: D = 6:1. Thus, for holes 165 mm (61/2 in.) diam, a charge 165 mm (61/2 in.) diam, and 990 mm (39 in.) in length would constitute a spherical charge. Crater Testing Crater testing for underground application is nor¬mally accomplished by drilling smaller than production sized horizontal holes into drift walls, placing relatively small charges of explosives in the holes at various depths of burial, firing the charges, and measuring the results. The results are scaled to develop charge weight size and depth of burial parameters for production hole sizes. In theory, this procedure is valid and is appli¬cable for surface blasting; however, in underground applications where the craters are inverted, there is some question as to whether extrapolation of the test results is appropriate. As well as the fact that craters are in¬verted with resultant gravity stresses, testing is often done in a waste rock drift where the properties differ from those in ore rock. There is also some question as to whether a 51 or 76-mm (2 or 3-in.) diam charge behaves similarly to a 165-mm (61/2-in.) charge in a confined borehole. Crater testing at Strathcona involved drilling 76-mm (3-in.) diam holes at various depths ranging from 1 to 3 m (3 to 10 ft). The test holes were drilled in the wall of a predominately waste drift as well as in a stope wall. Charges were placed at various depths of burial, were confined, fired, and the results measured. Although the results were somewhat erratic, possibly because of the aforementioned reasons, they did provide guidelines for basic design for production application. MINING METHOD APPLICATION FOR PILLAR RECOVERY Following initial crater testing, assessment, and tech¬nical evaluation, it was decided to use this blasting tech¬nique in a production application, and to incorporate

Jan 1, 1982

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

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Jan 1, 1940

By J. Macassi M.

La plata es uno de los recursos metálicos más importantes de Centromin, cuya complejidad de sus características requiere de un proceso de caracterización, identificación y conocimiento del comportamiento de los minerales que lo contienen, frente a los procesos existentes y a las nuevas posibilidades técnicas propuestas. En ese sentido, el texto plantea una alternativa hidrometalúrgica de solución para la recuperación de plata contenida en menas de hierro del tipo óxido sulfato, considerándola como una etapa de pretratamiento con una solución básica de hidróxido de sodio, de forma que esta tecnología permite beneficiar menas de naturaleza refractaria que contienen significativas cantidades de metales preciosos a un costo más económico.

Aug 26, 1988

By M. W. Springett

During the 1970&apos;s geostatisticians were often accused of not paying sufficient attention to geologic features in the computerised generation of ore reserves. Unfortunately this charge was all -too-often justified and the results of such studies tended to be oversimplified and failed to reveal the full potential of computerised reserve estimation and geostatistical methods. There were some notable exceptions to this trend. Techniques were described that paid a great deal of respect to the geologic characteristics peculiar to certain deposits (Rendu J-M, 1984, Barnes T. E., 1982). At Gold Fields several techniques have been developed for handling the geologic characteristics of ore bodies for reserve estimation purposes. These have developed over time and the evolution of their adoption has been a reflection of our own learning experience from the practical application of computerised reserve estimation on many precious metal s deposits. Problems Typical problems that have been experienced by ourselves and which we have learned to guard against and watch for are: ? "smearing" at geologic contacts ? confusion caused by disparate "selective mining unit sizes" and block sizes ? conflicting estimates generated by field staff using various versions of the polygonal method ? problems in generating meaningful variograms

Jan 1, 1985

By V. V. Lodeischikov

This paper covers some problems of treatment of complex gold ores and concentrates containing copper, lead, zinc, antimony and uranium. All these metals may present in ores both as major and associated components. The paper suggests recommendations for early determination of relative value of the metals in ores, which largely contribute to the selection of concentration and metallurgical flow sheets. The focus of this work is on treatment of complex ores in gold plants. The metal extraction technologies used by such gold plants are typically aimed at maximum gold recovery. Therefore, production of base metals as by-products requires significant updating of the existing process flow sheets based on economic, environmental and other considerations. At the same time, cooperation between gold plants and base-metal plants is of great importance. This paper provides an overview of worldwide existing and innovative processes of base metals extraction from complex and refractory gold ores.

Jan 1, 2014

By J. F. Fiset, N. Bouzoubaâ, N. Reynier, C. Laviolette, Canadian Nuclear Laboratories

Canadian Nuclear Laboratories (CNL) is developing a disposal strategy for radioactive wastes, which contain significant amounts of uranium, cesium, and mercury. There are safety and economic incentives for the extraction of these elements before permanent storage in deep underground repositories. In addition, the actinide elements are also an important component in rare earth element (REE) ores and considered as impurities. Actinides separation from REE tailings is a significant concern in order to manage radioactive nuclides. Reprocessing mine wastes may present significant financial and environmental benefits. Experiments were carried out with radioactive cemented wastes (RCW) prepared at CanmetMINING. For all of the tested RCW batches, solubilisation yields are above 97% for Cs and 98% for U and Hg. Sulphuric acid leaching with potassium iodide improved the solubilisation of Hg by oxidation and formation of the tetraiodide complex (e.g. HgI4 2-, Kf = 2.1030). Cs and Hg were removed selectively by using potassium cobalt hexacyanoferrate (KCFC) and Lewatit TP214 ion-exchange resin respectively. U recovery using the chelating resin Lewatit TP260 was the most efficient. The presence of high iodine concentration improved the U uptake by the resin. The best U elution was obtained using 1 M Na2CO3, which resulted in 99% U recovery. U was then precipitated as sodium diuranate. A counter current effluent recirculation was performed successfully on RCW. The process was also applied to submerged tailings collected from two U mines and an REE ore. This paper presents an innovative process, using sulphuric acid and potassium iodide, for U, Cs, a

Jan 1, 2016

By Pedro Lozada, Juan Prado

El objetivo que hoy se les exige a las técnicas de voladuras controladas es el de minimizar el daño al macizo rocoso remanente protegiendo la integridad de las paredes de la mina y asegurando a la vez la calidad de la fragmentación del material volado facilitando su extracción. Encontrar este equilibrio resulta estar entre los tres principales desafíos que enfrentan las grandes minas a cielo abierto, ya que deben lograr la estabilidad de la mina, dando cumplimiento al diseño de taludes, manteniendo al mismo tiempo buenos resultados en términos de fragmentación y productividad de sus equipos de carguío.

Sep 30, 2022

By C. Hilgendorf, J. Gingerich, J. R. Boniwell, M. Drouin, D. David on, J. Boldy

The initial discovery of gold in the Eastmain River area, 310 km northeast of Chibougamau, was made by Placer Development Limited in 1970. This discovery resulted from the drill testing of an airborne E .M. Target located in an area of extensive glacial overburden, within an unexplored Archean greens tone belt considered prospective for base metal deposits. Intermittent exploration by another company between 1973-1977 failed to obtain promising results and the existing claims were allowed to lapse. In 1981, Placer reactivated the project and initiated intensive exploration during the next two years which included aerial E .M. surveys, claim staking, ground E .M. and magnetic surveys, geological mapping, diamond drilling and attendant borehole P,E.M. surveys. Beginning in 1983, exploration has been conducted on a joint venture basis with Eldor Resources Limited, The greenstone belt under investigation has an arcuate distribution pattern relative to a centrally located granitic core. Occurrences of gold mineralization defined to date are located to the south of the granite along a strike length of 7 km, and appear to be restricted to the lowermost of two mafic volcanic cycles. This lower cycle is dominated by basaltic flows with subordinate rhyolitic flows and pyroclas ties. Concordant, variably magnetic, massive to talcose ultramafic units occur towards the stratigraphic top of the lower mafic volcanic cycle and appear to crudely define areas considered prospective for gold on a regional bases. Economic interest in centered on a s tratabound, siliceous, sulphide-bearing unit hosted in a 15- 25 metre thick sequence of rhyolite tuffs, basalts and intercalated ultramafics, The siliceous ore bearing unit ( interpreted as a recrystallized chert) contains from 10%-30% pyrrhotite and pyrite with subordinate chalcopyri te and minor sphaleri te. Alteration is characterized by mica development, silicification and carbonatization. Garnets are also occasionally present in close proximity to the ore horizon. The sequence is south facing, and overturned to the northeast at 35°-55°. Explora tory drilling in the vicinity of two conductive targets has been on 100 metre centres along a 1 km strike length of the prospective sequence. To the end of 1983, two ore zones (A and B) have been partially drill tested with estimated indicated and inferred geological reserves of 1 million tonnes grading 11.7 g/t gold , 14.7 g/t silver and 0.26% copper. The system is open at depth. This discovery is the result of the effective integration of geophysics and geology with contributions by many over a long period of time. Key events that can be emphasized are: ( 1) initial area selection of a remote greenstone belt in 1969 - albeit for base metal deposits , (2) utilization of an airborne E .M. system which successfully targeted the A Zone in 1970 , (3) the timely reactivation of the project in 1981 and the subsequent recognition that the original discovery represented a sulphide bearing stratabound horizon with implicit volcanogenic connotations , (4) the intensive effort in 1982-83 deploying aerial and ground geophysics coupled to volcanogenic concepts of ore distribution which resulted in the discovery of the B and C Zones in 1982, and (5) the integration of downhole P . E.M. surveys and geology to successfully delineate ore at increasing depths . To the end of 1983 , $2.6 million has been expended, and it is hoped that continued effort will result in the eventual delineation of an economically viable ore deposit for the joint venture partners .

Jan 1, 1984