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By K. J. Brouwer, J. P. Haile

"Construction of the first stage of the tailings disposal facility for the Montana Tunnels Project was completed in October 1986. The project is located 37 km south of Helena, Montana, and involves mining of42 million tons of weathered diatreme are at a rate or 13 500 tpd to produce a gold/silver dore, as well as lead and zinc concentrates. The tailings disposal facility for the project includes a fully drained tailings impoundment for storage of the tailings solids, and a process waterpond for storage of all decanted water prior to recyling to the mill. The over-all facility is designed to achieve the basic objectives ofminimizing seepage to the environment in the short- and longterm, and achieving a fully drained stable tailings mass suitable for immediate reclamation on completion of mining. Specific features of the design include a soil bentonite seal and drainage system within the tailings basin, a free draining embankment with a design flow-through capacity of 4000 US gpm and an ultimate height of 230 ft (70 m), and a soil bentonite lined make-up water pond with a capacity of 420 acre-feet (530 000 m'). The tailings will be discharged in a rotational sequence through multiple spigot offtakes to achieve a thin-layer low energy laminar flow. This 'sub-aerial' deposition results in higher stored densities and maximizes liquid-solid separation on the tailings beaches. The over-all design played a major role in the successful permitting of the project, and was significantly cheaper than a conventional wet disposal system, with overall unit costs for tailings disposal of less than US$0.36 per ton.This paper outlines the site characteristics and the tailings material properties and their influence on the design. Extensive use was made of naturally occurring free draining materials as well as bentonite modified soil liners within the tailings basin, embankment and retention ponds. General construction procedures for the various components are described, as well as on-going construction requirements utilizing materials from the open pit. Finally, the over-all costs and unit costs for tailings disposal are presented."

Jan 1, 1988

By James H. Courtright, Kenyon Richard

TOQUEPALA is a porphyry copper deposit in which mineralization is localized by a large breccia pipe formed in close genetic relation to intrusive rocks. The deposit is in southern Peru, 55 airline miles north of the small city of Tacna and the same distance inland from the port of 110. Quellaveco and Cuajone, geologically similar deposits, lie 12 and 19 miles north of Toquepala. Chuquicamata is 400 miles to the south. The deposit is high on the southwestern slope about 20 miles from the crest of the Cordillera Occidental of the Andes Chain. It lies in a mountainous desert where the steep southwesterly slope of the Andes is dissected by a succession of rapidly downcutting, deep canyons. Local topography is moderately rugged with a dendritic drainage pattern and an elevation of 8000 to 14,000 ft. Volcanic peaks along the crest of the Cordillera rise over 19,000 ft. Local precipitation, including a little snow, amounts to about 10 in. during January and February, but general runoff in the region is slight. Throughout southern Peru the springs and streams are widely separated. Crude canals irrigate small farms on terraced slopes along the streams and provide sparse subsistence to the semi-nomadic inhabitants. During the past decade, engineering and geological explorations of the region, as well as the mineral deposits themselves, have required construction of a network of several hundred miles of roads. Before this, roads extended only a few miles inland. Many areas still can be reached only by trail. Toquepala was briefly described in 19th century geographical literature as a copper deposit, and it received desultory attention from Chilean prospectors early in the present century. It was first recognized as a mineralized zone of possible real importance by geologist O.C. Schmedeman during an exploration trip for Cerro de Paso Copper Corp. in 1937. The discovery was late as compared to earlier recognition of Chuquicamata, Potrerillos, and Braden of Chile and Cerro Verde of southern Peru. This was due partly to the region's difficult accessibility but principally to the obscure character of the outcrop evidence of copper. From 1938 until 1942 Cerro de Pasco Copper Corp. partially explored the deposit by adits and diamond drillholes. This campaign was supplied by a 60-mule pack train continuously shuttling over a 30-mile trail. Northern Peru Mining & Smelting Co., a wholly owned subsidiary of American Smelting & Refining Co., undertook regional engineering stud- ies in 1945 and drill exploration in 1949. According to published data1 the deposit contains 400 million tons of open pit ore averaging a little over 1 pct Cu. It is currently undergoing large-scale development by Southern Peru Copper Corp., which is owned by American Smelting & Refining, Phelps Dodge, Cerro de Pasco, and Newmont Mining. Summary of Geology: The deposit is situated in a terrane composed of Mesozoic(?) and Tertiary volcanic rocks intruded by dioritic apophyses of the Andean Batholith. These formations are exposed in a northwesterly trending belt about 15 miles wide. Along the northeast they are unconformably overlain by Plio-Pleistocene pyroclastic rocks, which occupy much of the crest of the Andes, and along the southwest they are covered by the Moquegua formation of Pliocene(?) age. The mineralized area, oblong in shape and about 2 miles long, has been a locus of intense igneous activity. Several small intrusive bodies having irregular forms occur within and adjacent to a centrally located, large breccia pipe. The mushroom-shaped orebody consists of a flat-lying enriched zone of predominant chalcocite with a stem-like extension of hypogene chalcopyrite ore in depth within and around the pipe. This breccia pipe is relatively large and has been formed by repeated episodes of brecciation. Small satellitic pipes occur at random within a 2-mile radius of this central pipe. These too were individual sourceways of mineralization, although not always of ore grade. Within and around the zone of breccia pipes and mineralization there are a few faults and veins, but these are discontinuous random structures of minor significance. There are no regional or local systems of faults or other planar structures recognized which could account either for the mechanical development of the breccia pipes or for their localization as a group or as individuals. Hydrothermal alteration is pervasive in the zone of mineralization. Clay minerals appear to be abundant in places, but their percentages are undetermined. Quartz and sericite are the principal alteration products, and in many instances original rock textures are obliterated. The principal sulfides, hypogene pyrite and chalcopyrite and supergene chalcocite, occur mainly as vug fillings in the breccia and as small discrete grains scattered through all the altered rocks. Sulfide veinlets are relatively scarce. Sulfides are more abundant and alteration is more intense in certain rock units, such as the diorite and most of the breccias. Although the Toquepala mineral deposit is similar in most respects to the porphyry copper deposits of southwestern U. S., it most closely resembles the Braden deposit of Chile, as described by Lindgren

Jan 1, 1959

By Gary Hawthorn

"Laboratory testing, by Lakefield Research determined that, in spite of the relatively high pyrite content of the Golden Rose (GR) orebody, it would respond well to gravity concentration followed by either direct cyanidation or cyanidation of a flotation concentrate.Economics favoured the former choice which became the adopted process flowsheet.During the latter stage of the feasibility study, the Dekalb 500 tpd copper flotation mill became available in the Highland Valley of British Columbia. This plant was purchased by Emerald Lake Resources, and was moved to the Golden Rose minesite.The crushing and grinding circuit were retained and the original flotation circuit was replaced with a 400 mtd gravity plus cyanidation circuit. The plant commissioning commenced in August 1987, and by early September was processing 300 tpd.HistoryThe Golden Rose Mine was operated by Cominco during the period 1935 to 1941 processing 146000 tons grading DAD oz/t Au at a rate of 110 tpd .The process flowsheet consisted of gravity plus amalgamation followed by cyanidation.Metallurgical TestingThe underground mining areas were below the level of Emerald Lake, and had flooded on abandonment in 1942.Samples for metallurgical testing were collected from exposed faces underground once the mine had been dewateredin 1984.In spite of the anticipated high pyrite content of the ore samples, it was apparent from the success of the previous operation that the pyrite did not interfere with the gold recovery.Laboratory testing, at Lakefield, determined that, from a technical perspective, two metallurgical options existed, as shown in Tables I and 2.Note that the very high pyrite content of the ore precluded the possible direct marketing of concentrate. The process flowsheet, therefore, had to contain a cyanidation circuit for the operation to succeed financially.Although the anticipated metallurgy was excellent with either option, the 2.1 % increase in gold recovery with the gravity plus cyanidation option made this the clear choice."

Jan 1, 1989

By A. R. Hendrickson, C. F. Smith

Hydrofluoric-hydrochloric acid mixtures have been successfully used to stimulate sandstone reservoirs for a number of years. Hydrofluoric acid (HF) has a specific reactivity with silica which makes it more effective than HCl for use in sandstone. Kinetics of the reactions of HF have been studied to determine the related effects of reservoir composition, temperature, acid concentration and pressure on the spending rate of HF. Secondary effects from by-product formation are noted and described. Predictions are made concerning the improvement in productivity resulting from HF treatment of skin damage. The kinetic order of HF reaction in sandstone was experimentally determined to be first order, i.e., the reaction rate is proportional to concentration. HF reacts faster on calcite than on clay, which, in turn, is faster than the reaction rate of HF on sand. Static conditions retard the HF reaction rate. As HF is forced into cores, there is a temporary reduction as a function of flow rate and acid concentration. Extensive deposition of calcium fluoride in acidized cores was not observed. Although some CaF, was defected, it was not considered a major source of damage in cores containing moderate amounts of carbonate. Other fluosilicates could be potentially more dangerous than CaF, in reducing permeabiliry. INTRODUCTION Hydrofluoric acid has been widely used in stimulation treatments since 1935, when mud acid was introduced to the petroleum industry. Originally, this hydrochloric-hydrofluoric acid mixture was intended to remove mud filter cake, but it has since been successfully applied to many other oilfield problems. Mud acid treatments have been unusually successful in sandstone reservoirs where hydrochloric acid is unreactive due to a lack of enough calcite in the formation. The relatively small amount of hydrofluoric acid present (2.1 per cent) reacts with sand grains, clays and traces of calcite which are generally present in sandstone reservoirs. Since hydrofluoric acid (HF) is the key to mud acid success, this research effort has been dedicated to gaining a more thorough understanding of the basic chemical and physical principles involved as HF reacts. Hydrofluoric acid's reactivity with silica makes it unique in application. Other mineral acids such as hydrochloric, sulfuric or nitric are unreactive with most silicious materials which comprise sandstone formations. A typical sandstone reservoir may contain 50 to 85 per cent silicon dioxide, more commonly called sand or quartz. Hydrofluoric acid reacts as follows: 4HF + SiO2 + SiFO + 2H2O The silicon tetrafluoride (SiF,) is a soluble gas, in some ways similar to CO2, and is capable of undergoing further reaction when held in solution by pressure. These reactions will be considered in detail later. Kinetics of the reactions of HF have been studied to determine the effect of reservoir composition, temperature and pressure on the spending of the acid. Secondary effects from by-product formation have been noted and described. The individual reactions of HF on quartz, glass and clay are reported. Mathematical correlations have been drawn, then applied to studies of HF spending in cores obtained from actual producing sandstone formations. The research reported herein is only the beginning of a continuing approach to better understanding and use of HF in petroleum reservoirs. THEORY AND DEFINITIONS Through the years, a concentrated effort has been made to understand the effects of many variables on hydrochloric acid (HCI) spending in limestone. Hendrickson el al., have given mathematical relationships for HC1 reactions which made possible the engineered approach to acidizing. The same variables—temperature, acid concentration, formation composition, pressure and permeability-porosity relationships—which affect HC1 behavior in limestone also govern HF behavior in sandstone. Insoluble by-products of HF reaction have been isolated and identified. Their effect on fluid flow has been measured under varying conditions in an attempt to evaluate the extent of possible damage and means of eliminating it. In general, HF follows the same reaction paths as HCI. It will react with limestone and dolomite with speed and ease. Thin sections of acidized cores show the reaction of HF with limestone or calcite faster than its reaction with either clay or sand. When HF reacts with calcite (CaCO,), theoretically, calcium fluoride (CaF2) is precipitated, and has been blamed as a major cause of reduced permeability. On the other hand, pH and pressure such as that encountered in an underground formation under acid treatment definitely retard CaF2 formation,' so the whole question of CaF, deposition in wells is a subject for study.

Jan 1, 1966

By Richard Thomas Newcomb

Chapters 5a, 5b, and 6 discuss the long- run supply of minerals and the characteristics of reserve search and production peculiar to extractive industries. It is now necessary to complete the picture by examining supply and demand in commodity markets to show how mineral prices are determined and how both price and production trends might be predicted over long periods of time. These practical interests can be served best by reviewing first some of the characteristics of the short-run static market models familiar to all readers of elementary economic texts, and then extending the theory to models of a more dynamic nature. It is important at the outset to note that economic theory provides a wide variety of economic models based on varying assumptions concerning the behavior of economic agents, the properties of the goods, the time period considered, and so on. Each variety of model offers certain analytical advantages and suffers certain shortcomings. The art of the applied economist or practioner lies in knowing enough both about the technology and practice of his problem and about theory to select the appropriate models for his analysis. In our approach it will be useful first to consider the role of markets for minerals and mineral products in relation to the efficient determination of what and how goods are produced in the whole economy, i.e., a general system of markets, before getting down to the task of analyzing supply and demand conditions in specific markets. The former study is called general equilibrium analysis, while the latter is called partial. By equilibrium is meant a state of the sys- tem under study in which supply equals demand in each market, and the prices which equilibrate the quantities offered and asked are called equilibrium prices. The tendency of prices to return toward equilibrium values if temporarily displaced, or to converge to new equilibrium values if important new conditions for supply or for demand are placed, defines "stable" equilibrium systems. The short-run describes a period in which adjustments to fixed capital are not made, although variable factors such as labor or materials can be adjusted. The long-run envisions a period in which fixed capital and other such constraints can be adjusted. Thus, it may take a sharp increase in price to bring out additional supply of a factor in the short-run with existing plant, while new capacity might permit additional quantities in the long-run without an increase in supply price.

Jan 1, 1976

By D. Turnbull, R. E. Cech

FISHER, Hollomon, and Turnbull have developed a theory for the nucleation of martensite. They first tested the theory on Fe-C alloys and low alloy steels. The major factor influencing nucleation of martensite was considered to be statistical composition fluctuations occurring in small regions at high temperature and frozen-in on quenching. These local regions of varying size and composition serve as nucleation centers. They become supercritical, one by one, as temperature is progressively lowered, resulting in temperature-dependent or athermal transformation. Fisher next applied nucleation theory to substi-tutional solid solution alloys. Detailed predictions were made for Fe-Ni alloys because of the availability of free energy data on this system. It was shown that composition fluctuations that were significant energy-wise did not occur, and nucleation frequencies could be calculated from average properties of the system. Nucleation was predicted to occur as time-dependent and having the functional relationship to give a C curve of nucleation frequency vs temperature. The analysis further predicted that the nucleation frequency was extremely sensitive to composition. Experimentally, it would be found that the transformation in some compositions is so slow that measurable amounts will not form in a reasonable length of time. With other compositions, only slightly different, the nucleation frequency becomes so great that the material becomes transformed while still distant in temperature from the maximum nucleation frequency. On quenching an alloy of such composition, the observed transformation kinetics would be similar to those found in Fe-C alloys. Cech and Hollomon repeated experiments of Kurdjumov n which the kinetics of transformation were similar to those predicted by Fisher for Fe-Ni alloys. The alloy studied in this investigation contained 73.3 pct Fe, 23.0 pct Ni, and 3.7 pct Mn. Fisher,' using an idealized model for the extent of transformation as a function of the number of martensite crystals per grain of parent phase, derived nucleation frequencies from the transformation curves of Cech and Hollomon. Complicating influences such as coupling effects between grains in the polycrystalline specimens were neglected. Nevertheless, excellent agreement was found between the theoretically and experimentally derived nucleation frequencies. These experiments, however, could not provide a critical test of the theory. Experimental nucleation frequencies could vary widely from those calculated, depending upon the extent of deviation from ideal partitioning and the extent of coupling effects. Further, since the compositions of material theoretically analyzed and experimentally determined were different, the free energy changes involved in the experimental work could only be estimated. Also, the effect of heterogeneities on the transformation kinetics was not considered. For these reasons, it was decided that experiments designed to test the validity of the Fisher analysis must be conducted on binary Fe-Ni alloys, which were the ones considered theoretically by Fisher. The martensite transformation in Fe-Ni alloys has been the subject of considerable study. Machlin and Cohen have shown that transformation proceeds in a manner quite unlike that in any other ferrous alloy system. They found that single crystals would transform to a large extent in a single burst. In large grain polycrystalline specimens, frequently more than one grain and sometimes the whole specimen would transform at the same instant in this manner. Results on filings indicated that different particles would undergo the burst transformation at widely different temperatures. These results support the conclusion that the transformation behavior could not be described by a single nucleation frequency as would be the case if the nucleation were homogeneous. It appeared that further work was necessary to define the factors responsible for burst-type transformation, so that the conditions could be altered to favor homogeneous nucleation of martensite if such could be accomplished. This report summarizes the results of some experiments conducted with powdered Fe-Ni alloys for this purpose and the re-

Jan 1, 1957

By Dr. O’Neil Thomas J., Donald W. Gentry

There are mines that make us happy, There are mines that make us blue, There are mines that steal away the tear-drops As the sunbeams steal away the dew. There are mines that have lost the ore chute faulted, When the ore's forever lost to view, But the mines that fill my heart with sunshine, Are the mines that I sold to you. -Anonymous INTRODUCTION The preceding chapters have discussed, in various degrees of detail, the primary considerations necessary to perform a cash flow analysis of a mining-related investment opportunity. Although the previous discussions regarding the manipulations, calculations, and interrelationships of the various components of the pro forma income statement were illustrated in the text, many individuals are un- familiar with the format and appearance of a final spread sheet for the evaluation of a mining property. As a result, it seems appropriate to conclude this book with an abbreviated case study of an actual mine property (currently in production) showing the cash flow analysis and some back-up data. Obviously all of the supporting data and calculations for the case study presented cannot be provided; however, reference is made to the major assumptions utilized and to the appropriate procedures employed in the calculations, as described elsewhere in the text. The problem associated with using any case study for illustrative purposes is that one tends to be overwhelmed by detail. For instance, specific costs, prices, expenditure timings, and other variable values may be challenged because they either do not conform to one's expectations or are not consistent with specific operating experiences. When this occurs the instructional value of the problem becomes obscured. The purpose of most case studies, and particularly the one presented in this chapter, is to expose the student to the methodology, procedure, and format associated with performing a cash flow analysis of a mining property. It is not to laboriously wade through the specific details associated with operating and capital cost development, mine plans, or equipment schedules. It should be noted, how- ever, that the values represented in this case study result from specific mine plans, operating procedures, equipment spreads, manpower requirements, and other en-

Jan 1, 1984

By R. E. G. Caines, P. Alport

"Operating heavy equipment in Labrador, where the winters are 8 months long and the temperature frequently drops below -40°F (-40°C) for days and sometimes weeks presents unique problems in both the operation and maintenance junctions. Winterization programs must be well designed and implemented at the proper time. These programs must include the proper selection of fuel oil, engine oil, hydraulic oils, greases, cooling system fluids, etc. More often than not, modifications have to be made to the original equipment in order that it can be operated safely and economically in the Canadian North. Manufacturers usually design equipment for a less harsh environment and the user makes adaptations by trial and error.Operators have to be trained for the proper operation on ice- and snow-covered roads in an open-pit mine. Special emphasis must be put on haulage road conditions. In this paper, the authors will cover some specific maintenance procedures, equipment modifications and operator training problems experienced by the Iron Ore Company of Canada in its open-pit operation in Labrador.IntroductionT he Iron Ore Company of Canada is engaged in open-pit mining , concentrating and pelletizing of iron ore. The company operates in two provinces, Quebec and Newfoundland (Fig . 1) . The Knob Lake operation "", based in Schefferville, Quebec, has mine sites located both in Labrador and Quebec . The Caro l Lake Project is situated in Labrador, 25 kilometres east o f the Quebec border at Labrador City. The Carol operation mines, concentrates and pelletizes it s o re before rail transportation to the seaport o f Sept-lies via the Quebec North Shore and Labrador Railway."

Jan 1, 1984

By David Sinclair, Philip B. Bucky

The dimensions and shapes of mine structures may at present be determined by (1) field experience, (2) structural calculations, and (3) barodynamic tests. § None of these, however, provide information as to the stresses, their direction and distribution in a mine structure. For the design of pillars, the placing of openings relative to working places such as gangways, haulage levels, etc., this information is desirable. Photoelastic methods provide this and, while relatively new, their application in all engineering fields is assuming increasing importance. This paper, therefore, will present the method with a few applications to mining, and with sufficient reference for further study by the reader. PART I The Photoelastic Method The photoelastic method of stress measurement determines the stress distribution in a two-dimensional model by optical means. It makes use of the fact that when a piece of isotropic transparent material, such as glass, celluloid, or Bakelite, is stressed and viewed in polarized** light a picture is seen which indicates by its color bands of light, or fringes, the magnitude of the stresses, their direction and their distribution. General Explanation The stress at any point 0 in a structure (Fig. I), may be resolved into two components, P and Q, each a pure tension or compression, at right angles to each other, called principal stresses.' To fully identify the stress, the values of P and Q and their directions or angle F must be known.

Jan 1, 1940

By J. A. Brierley, Roshan B. Bhappu, D. H. Reynolds, P. H. Johnson

Although the economic importance of recovering copper by leaching of mine work dumps from open-pit operations has been realized for some time, serious attempts to understand and to improve such operations through scientifi research has been made only in recent years. This progress report covers the results of chemical, physical, and biological studies pertaining to the dump-leaching process and discusses salient features and parameters that may be helpful in improving such operations. The dump leaching of low-grade copper ores, as an integral part of the open-pit mining operations in the Southwest, has been practiced for the last 50 years and is increasing in importance as one of the major sources of copper. The recent acceleration in dump leaching can be attributed to: (1) the greater tonnages of low-grade ores mined and dumped at newer mines as well as resulting from increased stripping ratios at the older mines, (2) the relatively small investment required for dump leaching facilities per pound of copper recovered contrasted with expansion of milling facilities to recover equal amounts of the metal by raising tonnage throughout, (3) low labor requirements needed for leaching, and (4) the simple and continuous nature of the process needing little close supervision of trained operators. In the past, the mining companies have not worried about efficiency in dump leaching but have been satisfied with whatever metal recoveries could be made without much effort. However, because of the huge amounts of metal accumulating in the low-grade waste dumps, compared to the relatively small amount recovered annually, an operator today cannot leave dump leaching to chance. What appears to be a simple, empirical process is actually a very complex one involving a large number of critical factors that encompass Several Scientific disciplines, such as physics, chemistry, geology, biology, and hydrology. It is imperative that

Jan 1, 1970

By P. K. Chatterjee

"The operations of a large-scale open-cut mine are simulated to produce a user-oriented interactive digital model. SIMULA is chosen as the programming language and a sophisticated data base system is designed to consider the complex information network of large mines. The main features of the program are the mine planning and mine evaluation modules. These modules are responsible for the computation of truck-shovel performance and other equipment details in a mine. Stochastic methods are employed to generate event times. The use of a Data Base Management System and the data base provide an efficient method of information storage and retrieval. Truck allocation techniques have been examined in detail and comparisons between random dispatching and conventional methods are made. The model is intended for mine operators and engineers and the program is well structured and very readable. It is hoped that these features will make the model readily acceptable.IntroductionThe Department of Mining and Metallurgical Engineering at the University of Queensland has been conducting research into computer simulation methods in mining since 1972. The research has been mainly applied to actual operations, with both open-cut and underground mining being equally emphasized. The results of the research to date have not only strengthened the methodology of the approach to complex mining problems, but have also increased confidence in the use of modelling and computer simulation technique s in mining.The model described in this paper translates a complex LHD situation into a user-oriented interactive digital model using SIMULA. The present methods of evaluating truck dispatching and trolley assist are tedious and laborious. This model provide s the engineer with readily acceptable means of rapidly assessing the relative merits of these techniques in a large-scale open-pit operation. Poor utilization figures in many current operations (for example, 40% for trucks, 70% for shovels) are of serious concern to mine officials and computer-controlled dynamic truck dispatching systems have been considered by some of these mines to improve equipment utilization ."

Jan 1, 1981

By D. L. Madge

The text of this chapter was not received in time for printing.

Jan 1, 1979

By Charles Austin

IT has been known for several years1 that certain alloys of the Konal type, containing commercial cobalt (99.32 per cent C0 and 0.42 per cent Ni) and varying amounts of ferrotitanium, exhibit very high tensile strengths at 600°C. However, their pro-portional limits and yield points are rela-tively low. Published data1 for the alloys quenched in water from 950°C. and then tested at 600°C. are summarized in the upper section of Fig. I. Similar specimens that were aged at 650°C. for 72 hr., after quench and prior to testing at 600°C., gave the results shown in the lower section of Fig. I. TABLE I.-Analysis of Material I Percentages Constituent Sample Sample Sample 3125 3124 3123 Si 0.25 0.51 o.6o Fe 4.15 10.70 13.17 Ti I.15 3.58 4.70 Mn 0.30 0.41 0.42 Co (by difference)^ 94.15 84.80 81.71 The cobalt from all three samples, united and tested for nickel, gave 0.54 per cent Ni based on the weight of cobalt. In the present work the investigation of the properties of these materials has been extended by a study of the tensile deforma- tion characteristics in moderately long-time tests at 600°, 700° and 800°C. after quenching from various solution tempera-tures. In addition, some metallographic studies have been made to correlate with the deformation tests, and to obtain in- FIG. 1.-EFFECT OF VARYING AMOUNTS OF FERROTITANIUM (FE2TI) ON PHYSICAL PROPER-TIES OF COMMERCIAL COBALT. (C. R. Austin.1) Upper section. Tested at 600°C. after water-quenching from 950°C. Lower section. Tested at 600°C. after water-quenching from 950°C. and aging 72 hr. at 650°C. FIG. 2.-PORTION OF QUASI-BINARY DIAGRAM COBALT-FE2TI AS DETERMINED METALLOGRAPH-ICALLY. formation on the quasi-binary system Cobalt-Fe2Ti. The material, very kindly furnished by the Westinghouse Electric and Manufac-turing Co. in the form of hot-swaged rods of 3/16-in. diameter, analyzed as shown in Table I. The alloys, when originally made up,1 were based upon an iron-titanium com-pound, Fe3Ti, as indicated by the work of

Jan 1, 1940

By Louis A. Panek

This report deals with the combined friction and suspension effects in the reinforcement of horizontally bedded mine roof by bolting. Other reports have treated these two effects separately. Friction effect refers to the reduction of bending in a laminated roof due to the clamping action of tensioned bolts. If the beds tend to have unequal deflections, the bolts oppose this tendency, thus creating a suspension effect, which consists of a transfer of part of the weight of weaker or thinner beds to the flexurally more rigid beds within the bolted unit. The combined effects of friction and suspension on the bending of the beds, whether an increase or a decrease, are difficult to generalize because they depend on the bedding sequence and the bolt pre-tension. Some analysis of the bedding sequence usually is required to determine the critical bed within a set of bolted beds. However, if the suspension effect is substantial and if the bolts are pre-tensioned, then the greatest bending stress is likely to occur in the thickest bed. Findings are based partly on theory and partly on tests of mine-roof models in a centrifuge, which were found to verify the theory. Mathematical formulas are obtained for the change in maximum roof bending stress and for the change of roof deflection due to the combined friction and suspension effects. The maximum bending stress is needed for designing a roof-bolting system and the deflection for evaluating its effectiveness.

Jan 1, 1962

By John C. Loretto

"The paper identifies two of the main problems encountered in recent hydrometallurgical plants for the recovery of gold and silver: firstly, the plant recovery is lower than anticipated from the testwork; secondly, the metallurgical accounting figures cannot be reconciled.The principal reasons for the plant recovery being lower than the testwork are discussed, together with solutions that can minimize these problems both during the design stage and during the plant operation after the problems have manifest themselves.The reasons for the difficulties in metallurgical accounting are analyzed by reviewing the various components of the plant figures. Recommended solutions are given for minimizing this problem and the realities of obtaining accurate figures are discussed.IntroductionThe start-up over the last few years of new hydrometallurgical plants for the recovery of gold and silver has brought to the surface again a number of problems. Most of the problems have existed for some time and they manifest themselves mainly in two ways. Firstly, the plant results do not agree with the metallurgical testwork figures; secondly, the metallurgical accounting figures do not balance.Metallurgical TestworkThe plant results not agreeing with testwork figures are mainly due to:a) testwork samples not representative of the eventual plant feed;b) misinterpretation or misunderstanding of testwork ;c) plant losses;d) plant design not allowing testwork results to be achieved;e) operating deficiencies."

Jan 1, 1982

By R. Ricketts, D. G. Vanderveer, F. Kirby

"A program of mapping, sampling and testing to determine the potential of the aggregate resources throughout Newfoundland and Labrador was commenced in 1978. A 6-km-wide corridor parallel to all transportation routes was selected as the study area.The study consisted of a preliminary surficial interpretation followed by field mapping and sampling. Granular resources were sampled at a 500-m interval and non-granular materials were sampled at a 1.5- to 2-km interval. Data were gathered on landforms and stratigraphy as well as a field description of the sample and deposit types. Particle-size and lithologic analyses of the samples were also conducted.The locations of samples and deposits were mapped on 1:50,000 topographic base maps and summarized for publication at a 1:250,000 scale. Surficial geology maps were produced at a 1:50,000 scale. A computer program is being developed for the search and retrieval of field data.The first phase of the Inventory of Aggregate Resources program is nearing completion and to date has collected information on approximately 9,500 samples throughout Newfoundland and Labrador. This data will be used to provide f or the effective management of these vital non-renewable aggregate resources.The mineral aggregate industry in Newfoundland and Labrador is very competitive, but undercapitalized when compared to other provinces. The demand for aggregates is forecasted to increase to in excess of23 million tons per annum by the year 2002.In terms of over-all mineral production, the extraction of sand, gravel and stone has a significant impact when compared to other minerals in terms of value and tonnage produced.IntroductionThe province of Newfoundland has approximately 18,000 km of coastline. A large proportion of the population lives in coastal communities where, traditionally, beaches have been a major source of construction materials. Beaches have, therefore, often been subjected to overutilization, resulting in erosion of back beach areas or adjacent headlands."

Jan 1, 1982

By M. B. Mirza, F. D. Wright

Models of photoelastic material were made to simulate a mine roof that had cracked over the edge of the pillars and at the center of the span. Models were restrained from moving laterally outward so they acted as flat arches. Uniform loading was approximated by loading the models along the bottom at eight points spaced equally along the span. It was found that for models with depth to span ratios of 1:3 to 1:6, the maximum stress at the top of the center crack could be calculated directly by the following formula: Elastic stresses in mine roofs composed of sedimentary beds are usually computed by assuming that the immediate roof bed acts as a fully restrained beam uniformly loaded by its own weight and by overlying strata. Very little work has been done to analyze stresses in roof beds or beams which contain vertical (normal to the span) or nearly vertical cracks. Yet every mining engineer has walked safely under roofs that show such cracks. As far back as 1933, Bucky' made model studies of the effect of vertical cracks on the behavior of horizontal strata. He did not make any analysis of the stresses but did show that a cracked beam will stand safely under certain conditions, and also showed that a beam restrained at the ends can crack over the supports and in the middle without collapsing as long as the ends of the beam are not free to move laterally outward. The beam, or rather, part of the beam, acts like a flat arch and supports itself by pushing outward at its supports. In 1941 Evans2 treated analytically the problem of a beam cracked along so many planes perpendicular to the axis of the beam that no tensile stress can operate. He named this type of beam a "voussoir-beam" by analogy to the voussoir-arch. In 1962 ~a~cocks~ modified Evans's solution by including the effect of the shear stress component of the resultant force at the abutment. This paper gives the results of a photoelastic study of the stresses in a beam containing vertical cracks in the center of the span and over the supports. A

Jan 1, 1963

By Robert K. Koch

Laboratory-scale studies on defluorination of natural and synthetic fluorspars by pyrohydrolysis at 1,500° C. were made to determine the feasibility of developing a process for producing hydrogen fluoride from siliceous fluorspars. It was found that the reaction (at 1,500° C.) is not promising for process development, even though a silica-free offgas can be generated. The primary difficulty is the low hydrogen fluoride content of the offgases; an HF content of 6 percent was the best achieved. Previous research by the Bureau of Mines has demonstrated that offgases averaging 27 percent hydrogen fluoride can be produced under carefully controlled conditions at 1,280° C. The lower temperature process is a better method for production of HF. However, charges in which the CaF to Si02 mole ratio varied from 0.4 to 2 could be processed with similar results. Also, combined alumina and ferric oxide contents up to 25 mole-percent did not noticeably alter the rate of defluorination. Therefore, virtually any natural siliceous fluorspar would respond in essentially the same manner when exposed to water vapor at 1,500° C.

Jan 1, 1962

By A. C. Heath, E. H. Dix

THE importance of aluminum-silicon alloys in the light alloy field is now generally recognized. Where silicon was once considered detrimental to the properties of aluminum, useful alloys now contain as much as 13 per cent. Furthermore, silicon is present as an impurity in aluminum of even the purest grades. Iron is another important impurity in aluminum and like silicon is found in small quantities in the purest aluminum available today. The manner of occurrence of iron and silicon in commercially pure aluminum and many of its alloys exerts a very pronounced influence upon their properties. The effect of large quantities of iron in aluminum-silicon alloys is so detrimental that it has been found necessary to limit the allowable maximum percentage to as low a figure as is consistent with quantity production and satisfactory physical properties. The intimate relation existing between iron and silicon in aluminum and aluminum alloys makes necessary a more complete knowledge of this ternary system. This paper is divided into two principal parts, dealing respectively with aluminum-silicon solubility equilibria and the constitution of aluminum-iron-silicon alloys. PART I.-ALUMINUM-SILICON ALLOYS The early work on this system was largely confined to determining the general form of the diagram and in this respect the results are in good agreement. Briefly, aluminum and silicon form a simple eutectiferous series without compounds. Until recently the solid solubility of silicon in aluminum had not been accurately determined. Fraenkel1 (1908) stated that aluminum will retain 0.5 per cent. silicon in solid solution. Rosenhain, Archbutt and Hanson2 gave the maximum solubility as about 1.50 per cent. and added that "this solubility does not appear

Jan 1, 1928

By Ali M. Ghazi, Ronald J. Goble, Lowell S. Whiteside

"Geerite (Cu1.6S) is a rhombohedral copper sulphide which occurs naturally in epitactic replacement of sphalerite; cell dimensions closely approximate those of the host sphalerite.Geerite-type structures can be produced by leaching digenite (Cu1.8S) in ferric sulphate solutions. X -ray diffract ion and electron microprobe studies indicate that eight metastable geeritetype structures can be produced, depending upon the concentration of f erric sulphate solution used and the duration of the each. Although compositions cluster on or near values observed for natural phases, optical properties indicate that we are dealing with polymorphs of these phases. Leaching of chalcocite (CU2Sj in dilute f erric sulphate solutions produces similar results. In contrast, leaching of chalcocite in more concentrated solutions seems to produce true analogues of natural minerals.During froth flotation, sphalerite is activated with a monolayer of floatable copper sulphide with the composition, CuS, of the mineral covellite. However, covellite is not significantly more floatable than sphalerite, and has an incompatible structure because of the presence of covalently-bonded atoms of sulphur. Leaching ofdigenite was used to produce a polymorph of covellite with a geerite-type structure, compatible with that of sphalerite. This polymorph has a flotational behaviour similar to copper-activated sphalerite, suggesting that the monolayer has a geerite-type CuS structure. IntroductionIn an ideal close-packed sulphide structure the distance between sulphur layers, D, is equal to 2R 2/3, where R is the radius of the sulphur atoms. In sulphides with covalently-bonded sulphur layers, such as covellite (Fig. I), D will be an average of the number of covalentl y-bonded and close-packed sulphur layers, weighted as to the relative number of each type of layer. Both D and R are spacings which are readily identifiable structurally and on X-ray diffraction patterns. Goblel has shown that for copper-iron sulphides with metal-to-sulphur ratios between 0 and 2 and copper-to-total metal ratio s between 113 and I there are strong correlations between composition and measured D and R distances (Fig. 2). D can be represented by a single empirical equation relating interlayer spacing to composition; R must be represented by three equations relating the radius of sulphur atoms to composition. One of these equations (R, in Fig. 2) represents ideal close-packed behaviour, the other two represent sulphur radii greater than (R1) or less than (Rh) the ideal."

Jan 1, 1988