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By Flavio Begazo Mamani

Recientemente Southern Perú desarrolló un nuevo proyecto donde se modernizaba la planta de Fundicion y se reemplazaba el convencional horno reverbero por la tecnología del ISASMELT, capturando más del 92% de los gases generados durante el proceso de fusión de los concentrados de cobre. En ese sentido, el presente texto describo los efectos de la modernización de la Fundición sobre la calidad anódica. Antes de la modernización de la Fundición, los ánodos fueron producidos en la planta de ánodos ubicada en la Refinería de Cobre de Ilo. Ahora los ánodos son producidos directamente en la Fundición y como resultado, la composición química de los ánodos ha cambiado, principalmente en el contenido de bismuto, antimonio, plomo y zinc, siendo diferentes a los parámetros de diseño original de la Refinería. Asimismo, este factor impulsó a revisar el proceso de remoción de impurezas y los parámetros de la refinación electrolítica en la Refinería para evitar la contaminación del cobre refinado, logrando mantener la continuidad de la producción de cátodos de cobre de alta calidad superando los estándares interna-cionales de la norma ASTM-B115.

Sep 12, 2011

By Olga Frank-Kamenetskaya

The experimental results of study of atmospheric corrosion processes on Saint-Petersburg bronze monuments surface, received in situ and by means of simulating experiments, have been discussed. Influence of the time factor on corrosion film (patina) mineral composition has been analyzed. Direct dependence between patina mineral composition and environmental conditions has been proved. The sequence of patina phases crystallization has been established. Effect of the chemical composition and heterogeneity of artistic bronze and also monument surface structure on the rate of corrosion process has been regarded. The time variations of mass of the crystallized in the hydrochloric acid vapor phases have been investigated. The kinetic laws of their development at the initial stage of formation, indicating simple (non chain) reactions, have been revealed.

Jan 1, 2006

A major highlight of Fiscal Year 1993 is the completion of a new research facility dedicated to the advancement of robotics research and development. This new facility brings to the U.S. Bureau of Mines one of the largest and most unique facilities of its kind in the world. Located at the Bureau's Pittsburgh Research Center, the facility consolidates under one roof, all of the resource requirements of a first class automation and robotics research program. The Mining Robotics Development Center (MRDC) is the result of a major renovation of a large under-utilized warehouse-type structure and brings together the laboratories and offices previously housed in five separate buildings scattered throughout the Pittsburgh Research Center's 238-acre site. This new facility will support the computer-assisted mining research program, a high priority, multi-million dollar program that is expected to have a high impact on the health and safety of miners and coal productivity.

Jan 1, 1993

By W. R. Dengler

Dynatec Mining Limited contracted to sink a lS00 metre (4960 ft.) concrete lined shaft for Falconbridge Nickel Mines in Sudbury, Ontario, Canada. The shaft is 6.3 metres (20.S ft.) in diameter inside the concrete lining and is equipped with steel furnishings. Dynatec used the owner supplied 4.6 metre (l5 ft.) diameter double-drum hoist to sink the shaft with 12.S tonne buckets for muck removal and 4 cubic metre concrete buckets for concrete placement. One of the unique features of the shaft sinking operation is the use of a four-boom electric hydraulic drill jumbo permanently mounted in the shaft sinking stage. This paper will describe the performance obtained with the electric hydraulic drilling equipment and the high capacity muck removal and concreting equipment.

Jan 1, 1992

By Mike Amick

As virgin coal reserves dwindle, more and more mines are being opened in areas with previous mining. As addressed in this conference before, there are solutions available for designing mines extracting coal from reserves with known underground workings above or below the target seam. Often times, however, mining begins with limited knowledge of the underground workings. Even with careful planning, the proposed design has limited success because the ?facts? as presented during planning turn out not to be as accurate as one would like. At National Coal?s Mine #14, mining the Walnut Mountain seam, after driving the main entries in some 34 crosscuts, it was found during routine mine examinations that the pillars had separated from the roof in a few areas of the mains. Immediate roof conditions proved to be sound but questions arose concerning the stability of the main roof and floor. Core tests were done to better determine the rock strengths and characteristics. The paper presents the ideology used to tackle the effects of mining over old works in an operating coal mine. This paper discusses the analysis of the symptoms observed and the steps taken to ensure that this pillar separation was caused by the underlying old workings and that the separation was expected and controllable.

Jan 1, 2009

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

3.1 INTRODUCTION The final subsidence basin discussed so far is that developed long after mining has stopped. In a supercritical final subsidence basin, the central portion subsides uniformly. Thus any structures located there are not subjected to any final permanent surface deformation. But this statement is only true for the final subsidence basin. When the face is moving, the structures are subjected to dynamic deformations associated with the dynamic subsidence basin. For instance the house in front of a moving face (location 1 in Fig. 3.1) is subjected to the maximum tensile strain and the maximum positive curvahue. When it falls behind the face in location 2, it is subjected to the maximum compressive strain and the maximum negative curvature. When the house is far behind the face in location 3, it returns to its original condition with only uniform subsidence. However, the magnitude of the dynamic deformations may be sufficiently large to damage a structure in locations 1 and 2 before it reaches location 3. Therefore it is very important to study the characteristics of the dynamic subsidence basin. Under normal conditions, when the face has advanced for a distance approximately one-sixth to one-third of the seam depth from the setup entry, depending on the seam depth and the physical properties of the overburden strata, the movement of the over- burden strata reaches the surface, whereupon the surface also be- gins to deform. Thereafter, the surface point where the movement initiates is always located at a fixed distance ahead of the face. The angle between the vertical line at the faceline and the line connecting the movement initiation point on the surface and the face- line is the angle of advance influence, w. The angle of advance influence can be used to determine the distance, [ 1], which is the influence zone ahead of the face. w is approximately 20O to 35 O for the northern Appalachian coalfield (Peng and Geng, 1984). The movement of each surface point is a relatively complicated process both in time and in space, because the relative position between the faceline and each surface point varies continuously, so that the direction and rate of movement for each surface point differ. In the following sections, the movements of each surface point on the major cross section are illustrated. 3.2 DYNAMIC SURFACE MOVEMENTS Dynamic surface movements include dynamic subsidence and displacement. The movement of each point on the subsidence basin is different as the face advances from 1 to 6 (Fig. 3.2). Surface points A and B, being far in front of the moving face, subside and displace toward the face (i.e., from A to A' and from B to B'). Point C, which is quite near but in front of the moving face, subsides continuously but displaces initially toward the face and then turns to pure subsidence ending up at C'. Point D, which is behind the face above the gob, subsides continuously, but displaces initially toward the face when it is in front of the face and reverses to follow the moving face when it is behind the face, with the amount of displacement toward the face being larger than that following the face. As a result the net displacement is toward the center of the gob (i.e., from point D to D'). The same statement is true for point E. For point F, the net displacement is zero, i.e it appears to have dropped down vertically to F'. For points G to I, the net displacements are in the direction of face advance except the magnitudes are increasing in sequential order. Point 1, has reached the final dynamic position of 1'. In other words, the face has passed sufficiently away such that it does not affect the movement at point I. All other points (A to H) are still within the face influenced area. 3.3 DEVELOPMENT OF DYNAMIC SUBSIDENCE PROFILES When the face has moved for a distance of h/6-h/3 from the setup entry, surface subsidence reaches the surface. At this time the surface subsidence basin is very mild (S, in Fig. 3.3), and it is symmetrical about the center. As the face continues to advance, the maximum subsidence continues to increase, and its location moves with, but falls farther behind the face. When the face has advanced a sufficient distance, e.g., (0.9-2.2)h, away from the setup entry, the maximum subsidence reaches the maximum possible value (curve S,). At this time, the dynamic subsidence profile is fully developed. Thereafter both sides of the dynamic subsidence profile become stabilized. The subsidence profile on the face side moves with the advancing face in an orderly manner (curves S, and S,). When the face stops, the face side of the profile continues to subside and becomes stabilized (curves;) sometime after. Curve S; is the final subsidence profile (see Fig. 2.22). 3.4 DEVELOPMENT OF DYNAMIC DISPLACEMENT AND SLOPE PROFILES The dynamic displacement profile moves with the advancing face (Fig. 3.4), but changes according to some specific rules. When the face has advanced to A, the surface begins to move and its dynamic displacement profile has two maximum values (curve Ua): one positive near the setup entry moving toward the face advancing direction and the other negative on the face side also moving toward the face advancing direction. As the face continues to advance, the positive maximum displacement on the setup entry side increases and moves toward the face continuously. The same statement is true for the negative maximum displacement on the face side. When the opening increases to the critical size, both the maximum positive and negative displacements reach the peak values and remain constant thereafter. Notice that when the opening becomes critical in size, the point of zero displacement no longer moves with the advancing face. Rather it spreads into an area of zero displacement (curve Ue) and expands as the face advances. When the face stops, the displacement on the face side continues until it becomes stabilized (curve Úe). Curve Úe is the final displacement profile. The ratio of maximum dynamic to final displacement ranges from 60 to 75%. The dynamic slope profile follows the same patterns of change. The ratio of maximum dynamic to maximum final slope ranges from 50 to 80%. The maximum slope decreases with in- creasing rate of face advance (Peng and Geng, 1984). It must be pointed out that except for those near both sides of the profiles, each point on the displacement and slope profiles is subjected to a cycle of positive and negative displacements and slopes depending on its relative position with respect to the face.

Jan 1, 1992

The Annual General Meeting of the South African Institute of Mining and Metallurgy was held in Kelvin House on Wednesday, 23rd August, 1978. Mr P. A. von Wielligh (President) was in the chair and declared the meeting open at 16h00. Obituaries The President: It is my sad duty to announce the death of the following members of the Institute : T. K. Prentice and W. E. Gooday, Honorary Life Fellows and Past Presidents; B. S. Allen, J. A. Becking, L. H. Mellor, C. A. Pratt and W. T. Stobart, Life Fellows; J. P. Burls, E. Cameron, M. L. Fitzgerald, C. G. Hinds, A. J. Patrick, H. O. Reisener, N. D. Shuttleworth, N. Soldan and J. Whitehouse, Fellows; T. Gemmell, J. S. A. Oelofse, W. D. Rowe, R. L. J. Shearman and G. M. Thain, Associates. As a mark of respect to the memory of the deceased and in sympathy with the bereaved, I ask you to rise and observe a few moments' silence. Welcome The President: It is now my very pleasant duty to welcome our distinguished guests and visitors. We are pleased to have with us Mr D. A. Etheredge, Vice¬President of the Chamber of Mines of South Africa, and Mrs Etheredge. Mr Etheredge this afternoon represents the mining industry at our meeting. The importance of the close link that over the years has been maintained between the industry and our Institute speaks for itself. We are, therefore, grateful that he and Mrs Etheredge have managed to find the time to be with us this after¬noon.

Jan 1, 1979

By Richard A. Hawkins

O the American coal man, Australian coal mining most appear to have little, if any, influence on American coal-mining practice and to bear little relation to it. Actually, the relationship has been close, for Australian operators have carefully studied American practice by reading American technical publications, by consulting American engineers, and by visits to Ameriean operations. They have leaned heavily on America for their supply of mechanization equipment and many American mining-machinery manufacturers are active in that country, supplying its needs and furnishing technical assistance. Australia's influence on American coal mining has been obscure and probably negligible. However, it seems reasonable to assume that this influence will increase. The wartime development of air transportation will inevitably bring these two countries into much more intimate contact than existed in recent years. American servicemen. many of them former miners and coal-mining technicians, have visited Australian operation, while stationed in that country, have talked with operators and men over there, and undoubt-

Jan 1, 1945

By Hamid Maleki

A comprehensive study consisting of stress determinations, core logging, laboratory testing, and numerical analyses was conducted to investigate the cause and potential alternatives to the coal bump conditions that developed in the Second North Panel of the Little Dove mine near Buntington, Utah. Roof rock characteristics were such that a cave did not develop above this 420 ft wide panel. After 700 ft of pillar retreat, severe coal bump conditions developed during pillar mining. The coal bump conditions resulted from excessive pillar stress which was caused by load transfer from the extracted, noncaving, portion of the panel. Site specific data and numerical modeling techniques were used to back-analyze the bump prone condition. This back- analysis lead to the identification of critical stress, closure, and energy release levels. These critical levels were then used to evaluate alternative pillar mining sequences in the hope that the potential for coal bumps could be reduced. It was found that while alternative sequences offered some potential improvement over the pillar mining sequence used in the mine, alternative pillar mining sequences would not significantly reduce the potential for bumps. For the roof characteristics at this mine, coal bumps are best avoided by developing narrow panels or significantly wider panels, which incorporate properly sized barrier pillars. Narrow panels, on the order of 300 ft, would not cave, but panel pillar stresses and the potential for umps would be moderated by the reduced r' panel width. Wider panels, of approximately 650 ft, would allow a cave to develop, also reducing panel pillar stresses and coal bump potential.

Jan 1, 1987

By Kimio Itagaki, Torn Matsumoto, Makoto Ohtsuka, Kentarou Harada

"The structure and hydrogen-absorption properties of LaNis thin films were investigated in the present study. LaNis thin films were prepared by a sputtering method with a radio-frequency power, Ws, of 50 -400 Wat substrate temperature, Ts, of323 -573 K. Nickel content of the films decreased with increasing Ws, however, there was no dependence on Ts. At low Ws and low Ts, amorphous films were deposited. In other conditions, the films were polycrystalline. The films prepared at high Ws and low Ts, or at low Ws and high Ts, exhibited only a (h k · 0) peak in the X-ray diffraction profiles. This indicates that the c-axis of crystalline films was oriented parallel to the substrate plane. At high Ws and high Ts, the films with columnar structure were obtained. Hydrogen-absorption capacity of the thin films increased with increasing Ws. The pressure plateau did not appear on the pressure-composition isotherms. The films deposited at high Ts represented a higher equilibrium pressure.IntroductionThin films of hydrogen storage alloys can be applied for an electrode of rechargeable battery or fuel cell, a filter for H1 purification and various kind of sensor. Some studies on the preparation of the thin films of LaNi5 intermetallic compound, which is a typical hydrogen storage material, have been done hitherto. Adachi et al. , 2) and Sakaguchi et al. 3) prepared amorphous and crystalline LaNi5 thin films using flash evaporation of alloy powders and measured the pressure-composition isotherm for the prepared film utilizing a quartz crystal mass monitoring (QC:MM) method. It was found that the prepared amorphous film absorbed hydrogen without disintegrating into the powders and the hydrogen-absorption capacity of the films increased with the degree of crystallinity. Sakai et al. 4) reported that the maximum hydrogen storage capacities of amorphous and crystalline LaNi5 thin films, fabricated by a RF sputtering method using combination targets, were about 40 mA h g-1 (LaNisHo.7) and 160 mA h g-1 (LaNi5H2.6), respectively. Uchida et al. 5, 6) prepared LaNi5 thin films using flash evaporation of alloy powders and measured volumetrically the pressure-composition isotherm for the prepared film. It was reported that the hydrogen-absorption capacity of the films changed by the annealing treatment."

Jan 1, 1998

By Terkel Rosenqvist

THE desulphurization of molten iron and steel is a very complicated process. One way to arrive at a better understanding of this process is to break it down into several simpler chemical processes that can be studied individually in the laboratory. For a study of the different factors that influence the equilibrium distribution of sulphur between liquid metals and slags, several simpler equilibria may be investigated. One very important subject is the determination of the escaping tendency of sulphur in the liquid metal and its dependency on temperature and composition of the melt. Several papers in this field have recently been published.', ' Another subject is the study of the sulphur capacity of the slag. A molten slag is indeed complex, and even if sulphur distribution data for a large variety of molten slags may give empirical data about their desulphurizing power, the importance of the individual components is still not quite clear. It is accepted generally that lime is the most important desulphurizing component in the slag. The present investigation has as its purpose to study the desulphurizing power of lime in its standard state, and to provide a basis for thermodynamic calculations of the desulphurizing power of various lime-containing slags. The standard state of lime at steelmaking temperatures is solid calcium oxide, CaO. It can react with sulphur to form solid calcium sulphide, CaS. The relative stability of calcium oxide and calcium sulphide is expressed by the free energy of the reaction: 2Ca0 (s) + S1 (g) = 2CaS (s) + O2 (g) The existing free energy data for this reaction, listed by Kelley5 nd Osborn,' are uncertain to about 10 kcal and are of limited value for a calculation of equilibrium constants. Under the conditions prevailing in a melting furnace, the sulphur pressure may be expressed conveniently by the ratio H,S/H2 and the oxygen pressure by the ratio H,O/H, (or CO,/CO). The desulphurizing power of calcium oxide may, therefore, be studied by the reaction CaO + HIS = CaS + H2O. A study of this reaction may be complicated by certain side reactions: Water vapor and hydrogen sulphide may react. to form sulphur dioxide, and calcium sulphide may be oxidized to calcium sulphate. A thermodynamic calculation shows that these side reactions will be suppressed to insignificance if the equilibrium is studied in the presence of an excess of hydrogen. The apparatus used is shown in Fig. 1. About 10 g calcium oxide and 20 g calcium sulphide (laboratory qualities) were intimately mixed, and some water was added to make a thick paste. The paste was put into a thimble of zirconium silicate, which was placed within the constant temperature zone of a furnace, and capillary refractory tubes were attached in both ends. After the mixture had been heated in dry hydrogen at 1000°C for several hours all Ca(OH), and CaCO, had decomposed and CaSO, was reduced, so only CaO and CaS remained in the thimble forming a porous plug. The mixture was examined by X-ray diffraction after the initial reduction in dry hydrogen as well as after the subsequent experimental runs up to 1425 °C. It was shown that crystalline calcium oxide and calcium sulphide were always present together in about equal amounts. The unit cell edges were found to be 4.80A for CaO and 5.68A for CaS in good agreement with existing literature values." This shows that the mutual solid solubility is very small, and that the compounds are present in their standard states. Purified hydrogen was passed through water sat-urators kept at constant temperature in a thermostat bath. The amount of water vapor saturation was checked by means of a dew point method, not shown on Fig. 1. The gas mixture was passed through the capillary inlet into the furnace, where it was sifted through the porous plug of calcium oxide and calcium sulphide. The hydrogen sulphide present in the outgoing gas was absorbed in a zinc acetate solution and the hydrogen was collected over water. When one liter of hydrogen had been collected, the amount of hydrogen sulphide was determined by iodometric titration. As one molecule of H,O is used for the formation of each molecule of H,S, the equilibrium ratio H,S/H,O would be , where (H,O) is the molar concentration in the ingoing gas, and (H,S) the molar concentration in the outgoing gas. In the present work (H,S) was always very small compared to (H20). In order for the observed H,S/H20 ratio to represent the true equilibrium ratio the gas flow has to be: 1—Sufficiently slow to give a complete establishment of equilibrium, and 2—sufficiently fast to counteract thermal diffusion. Incomplete reaction would give a value decreasing with increasing flow rate, and thermal diffusion would give a value increasing with decreasing flow rate. When inlet and outlet tubes of about 2 sq mm cross-section were used, the observed gas ratio was independent of the flow rate between 15 and 125 cc per min, Fig. 2. In this range, therefore, the observed gas ratio represents true equilibrium.* For the rest of the in-

Jan 1, 1952

By Richard Havens

Acidic and basic uranium leach plant residues located on the Navajo Indian Reservation, Tuba City, Ariz., were successfully stabilized against wind erosion using a relatively low-cost chemical method. An elastomeric polymer chemical was applied to the dike areas and a calcium magnesium lignosulfonate to the beach areas of three tailings ponds. The water-soluble chemicals were applied with an automated sprinkling system. The stabilization cost was $335 per acre for the 34.5-acre tract.

Jan 1, 1969

By C. C. Wright, R. J. Day

Although data on the exact tonnage of recoverable coal silt are not known, the quantity produced in 1943 was estimated to be over five million tons for the anthracite region of Pennsylvania alone. Since that time the recovery of the No. 5 buckwheat size previously included in silt estimates has increased greatly thus reducing the total silt tonnage. As a compensating factor, however, the methods of recovering silt previously escaping into the rivers and streams have been markedly improved so that the net figure may not have changed materially. Comparable estimates for bituminous coal are lacking, but undoubtedly the silt represents a significant tonnage. With the trend toward increased wet cleaning of the smaller sizes of bituminous coal, it is to be expected that even greater quantities of silt will be available in the future. The raw silt from anthracite collieries usually contains a high mineral matter content, but technological developments in fine coal cleaning have been making rapid advances and it appears probable that more extensive use of these methods will be employed in the near future. Bituminous coal silt in many instances contains a similar high mineral matter content. As with anthracite silt, however, the cleaning of this material, where necessary, appears both technologically and economically feasible. With a view to upgrading the size and value of silt, considerable laboratory and pilot plant work has been conducted by a number of organizations on what is commonly known as pelletization. Actually the process is one of extrusion in which the moist silt is extruded from a die in the form of rods, which usually break into relatively short lengths—a few inches long—as they fall from the die. Fig I is a photograph of some extruded 3/8-in. diam pellets and of the anthracite silt from which they were produced This technique of producing larger sized material from fines is by no means new, having been used extensively by the industrial carbon and clay industries for years.' The application of the process to coal silt appears, however, to be relatively new. The first major program of investigation dates back only to July 1945, when a group of engineers representing the Anthracite Institute, The Pennsylvania State College and the J. F. Pritchard and Co. conducted preliminary tests on the extrusion of anthracite silt at the Floridin Co. plant at Warren, Pa. Since that time these three organiza-tions—individually, together, and in cooperation with other interested groups— have actively pursued work on various phases of the problem. Much of the early history and details of the laboratory and pilot plant work by the Anthracite Institute, has been reported by Mulcey and Eckerd.2 It is the purpose of this paper to report additional information on the process, especially with respect to the factors influencing the quality and use properties of the product, and the current develop-

Jan 1, 1949

By Daniel G. DiCanio

This report describes a computer-based model for estimating the tangible costs of injuries and deaths from work-related accidents in underground coal mines. The study undertook to identify and quantify the elements of accident costs, including field visits to coal mines, insurance carriers, and public agencies. The mathematical models developed for each tangible element of accident cost are presented in Volume I of the report. A listing of the computer programs which comprise the model and instructions for using the programs are presented in Volume II. Statistical distributions of mining family characteristics, data on 95 miners killed in underground accidents, and other basic data used to develop the model are presented in Volume III. Application of the model to a data base of 9,286 accidents which occured in underground bituminous coal mines during 1974 yielded a total of $56,900,000 in tangible costs to industry, mining families, and public agencies. Loss of income to the families of miners injured or killed in these accidents accounted for 47 percent of the total.

Jan 1, 1976

By Anderson Lúcio de Souza

Foram realizados testes completos de pelotamento para avaliar os efeitos da adição de uma solução de cal hidratada ao pellet feed na etapa de pelotamento. Foram preparadas soluções com diferentes concentrações de cal hidratada e após a mistura das mesmas no pellet feed foram realizados testes de pelotamento onde os principais parâmetros de qualidade das pelotas cruas geradas foram avaliados. O diâmetro médio final das pelotas foi o principal parâmetro influenciado pela adição de cal hidratada. A presença dos íons Ca2+ e CaOH+, em função da dosagem de cal, altera o estado de dispersão do minério de ferro, afetando assim o processo de formação das pelotas. Ainda, os cátions Ca2+ presentes no pellet feed interagem com o aglomerante orgânico e com a bentonita, reduzindo a capacidade desses agentes em controlar o crescimento das pelotas.

Aug 1, 2018

With the exhaustion of orebodies close to the main haulage shaft of the Emperor mine, an increased rate of development, aimed at earlier investigation of values disclosed by exploration drill holes has been attained with an overall reduction in unit costs.This paper describes the techniques used in rates of advance up to and exceeding 1,000 ft per 28-day period.

Jan 1, 1966

By Gordon Miner

GENERAL DESCRIPTION At the Star mine, Hecla Mining Co., located in Burke, Shoshone County, ID, the mining method is horizontal slicing (timbered or untimbered) cut-and-fill using hydraulic sand backfill. Back stoping as well as breast stoping is used; see Fig. 1. The annual produc¬tion from all cut-and-fill mining is 259 816 t (286,400 st). The average depth ranges from 366 to 2408 m (1200 to 7900 ft). A majority of the tonnage comes from an average depth of 2240 m (7350 ft). Average travel time to the stope is 1 hr. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size and Ground Conditions The ore body is 610 to 1219 m (2000 to 4000 ft) along the strike and 2408 m (7900 ft) on the dip, with an average vein width of 2.6 m (8.5 ft). The dip is essentially vertical. The ore body is heavily sheared and broken on the main fault structures; it is more competent on the wall veins but everywhere subject to rock bursting at depth. The hanging and footwalls are strongly fractured on the main fault structures, again more competent but still fractured on the wall veins. All host rock is Revett quartzite. Value and Production Requirements The ore value is $56 per ton. Production require¬ments are 1089 t/ d (1200 stpd), five days per week. Why Cut-and-Fill? The slusher extraction method is the most satisfac¬tory and practical for narrow vertical veins with sheared walls. The method must be quite selective to be com¬mercial. The veins are too narrow for rubber-tired ex¬traction and the grade is not high enough to make resuing practical. DETAILED DESCRIPTION Development Access to the main shaft is by the 3.2-km (2-mile) main haulage tunnel. The level intervals in the vertical timbered shafts are 61 m (200 ft). Vein development is by crosscut and laterals located 30.5 m (100 ft) into the vein wall. Vein crosscuts are on ±61 m (200 ft) centers. Access to the vein is by raises carried up with the stope. The raises are equipped with four wood¬cribbed chutes (two orepasses reinforced with angle irons, one manway, and one timber slide, Figs. 2-6). The initial cut is located 6.7 m (22 ft) above the main level (Fig. 1). Stoping Cycle Detail Sampling: Grab samples are taken from the mine cars. Channel sampling of the faces and backs is done when necessary. Drilling: For stope drilling, Ingersoll-Rand 300M percussion drills are used, with Ingersoll-Rand 300M and Gardner-Denver RB83M stopers, used for back stoping. Breasting down is done with Gardner-Denver S83FM jacklegs. The back stoping rounds are 0.9 to 1.8 m (3 to 6 ft) and the breast stoping rounds 1.5 to 3m (5 to 10 ft). Blasting: For stope blasting DuPont Tovex 100 watergel with fuses and igniter cord is used primarily. Approximately 60% of the blasting is done with am¬monium nitrate-fuel oil (ANFO). For development headings and shaft sinking, DuPont Drivex watergel with Ensign Bickford Nonel Primadet primers is is used. Mucking: Broken ore is removed entirely by slusher in the stopes. These are mostly 3.7-7.5-kW (5-10-hp) Ingersoll-Rand air-powered or 7.5-37.3-kW (10-50-hp) Ingersoll-Rand electric-powered. In the horizontal de¬velopment headings Eimco 12B and 21 and Atlas Copco LM-56 overshot mucking machines and a 0.76 (1 cu yd) hydroelectric Wagner HST-1 LHD (load-haul-dump) unit are used. Rockway for Extraction: The cribbed chutes (Fig. 7) are nearly vertical and hexagonal in shape and are made of spaced 152 x 203-mm (6 x 8-in.) timber, reinforced with varying weights and sizes of angle iron, depending on wear requirements. Rectangular cribbed chutes are used where ground pressure is not a problem. Cribbing is added in 2.4 or 3-m (8 or 10-ft) extensions accord¬ing to the floor height (whether timbered or untimbered). See Figs. 8 and 9. Backfilling: Backfilling is done with hydraulic sand consisting of classified mill tailings at 65% solids. The fill contains approximately 20% 325-mesh fines. The tailings are pumped 3353 m (11,000 ft) at 50-55% solids 22.7 dry t/h (25 dry stph) with a duplex piston pump. Cycloning increases the tailings density to 65% solids. Fill storage is in a 998-dry-t (1100 dry st) agi

Jan 1, 1982

By Dennis Ofiara, Michael Lewis

An experimental 1.75 meter Shaft Boring Machine was designed and manufactured by The Robbins Company. The drill was used by contractor Drillcon Raise AB to drill test shafts at an SKB underground test facility. SKB is the Swedish authority responsible for safe management of nuclear waste materials. This “SBM1.8” Shaft Boring Machine was used to drill shallow test shafts ingranitic rock at SKB’s Aspo underground laboratory. The drill used a unique vacuum mucking system. The bored shafts were held to very strict tolerances for bore diameter, straightness, and verticality. This drill was launched and operated from a mobile launcher. Headroom and working space were limited. The SBM1.8 drill produced acceptable test holes. SKB will carry out several studies to determine the suitability of this geology, and of this technique, for safe nuclear disposal. In addition to future use at SKB, many applications are foreseen, particularly in the civil construction industry.

Jan 1, 2005

By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By Robert L. Montgomery

Heats of solution of samarium (III) oxide and chloride in hydrochloric acid and of samarium (III) oxide, sulfate, and hydrated sulfate in sulfuric acid were measured. The heats of solution of this hygroscopic oxide are larger than those from the literature, probably because of more complete removal of water in the present study. From the literature value for the heat of combustion of samarium, the heats of formation of samarium (III) chloride and sulfate are -244.18 ± 0.41 and -928.5 ± 1.6 kcal./mole, respectively. The heat of hydration of the sulfate is -34.7 ± 1.0 kcal./mole. The heats of formation of other samarium com-pounds have been recalculated, using the new data.

Jan 1, 1959