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By M. C. Fuerstenau

A number of interesting and. important phenomena were revealed during a study of the response of beryl to sulfonate flotation (1). Perhaps the most important observation is the effect of various cations on flotation response. Experiments showed that in the vicinity of pH = 3. 0, about three ppm Fe.+++ increased recovery of leached beryl from about 60 to 100 per cent, whereas about four ppm Ca++ reduced the recovery from 60 to 2 per cent. With the exception of ferric iron, all of the other cations investigated reduced flotation recoveries drastically at this pH. As postulated in the paper "Sulfonate Flotation of Beryl", (1) even with low concentrations of polyvalent cations and sulfonate (on the order of one ppm), cation¬ sulfonates are precipitated immediately in solution. That is, blueish-white clouds .were noted to form after given additions of salts to. known solutions of sulfonate. Furthermore, it was also shown that one-hundred-fold more ferric iron than necessary to form the precipitated cloud had to be added to effect complete flotation. In other words, some form of solid ferric sulfonate was found to function as the collector in these systems. The form of the compound was suggested to be Fe (RSC3)2OH and that bonding to the surface occurs through the hydrogen of the hydroxyl. As the form of this compound was shown to be dependent on the extent of hydrolysis of ferric iron, it is to be expected that ferric iron will be unique in some pH region. Further, if the presence of hydroxyl in the neutral precipitated compound is of primary importance, then other cations should function similarly but at different values of pH, since they will hydrolyze at different concentrations of hydroxyl ion. To determine whether the premise of hydrolysis is correct, the study was extended to quartz systems lit the presence of various cations at various values of pH.

Jan 1, 1964

By H. I. Aaronson, C. Wells

THIS is a preliminary report on a detailed metal- lographic investigation of the bainite reaction in a plain carbon hypoeutectoid steel. In view of the current interest in the bainite reaction and of the controversies which still exist about even the basic phenomenology of this mode of transformation, it was considered desirable to present a summary of the more important results obtained to date prior to the completion of this investigation. The steel used contained 0.29 pct C, 0.76 pct Mn, 0.25 pct Si, 0.005 pct P, and 0.007 pct S. The alloy was homogenized for 48 hr at 1250°C. Individual specimens were then austenitized for 30 min at 1300°C and isothermally transformed for various times at temperatures above and within the bainite region, ranging from 775" through 500°C at intervals of 25°C. These specimens were carefully examined on a research metallograph; a few specimens of especial interest were also studied electron microscopically.' Bainite may be defined as the aggregate structure resulting from the precipitation of carbides from supersaturated ferrite which formed by nucleation and diffusion-limited growth. Two groups of experimental observations made during the present study provide support for this mechanism of the bainite reaction: 1—at 625°C, partial conversion of pro-eutectoid ferrite to bainite began after more than 99 pct of the austenite had decomposed and 2—the precipitation of bainitic carbides occurred only within the ferrite phase. The proeutectoid ferrite and the bainite reactions in the steel studied were found to be morphologically continuous. Following the classification scheme of Mehl and Dube,2 three principal morphologies of proeutectoid ferrite were identified in the temperature region (625" to 550°C) in which the two reactions blend—grain boundary allotriomorphs," * Grain boundary allotriomorphs are crystals which nucleate at and grow preferentially along matrix grain boundaries. Widmanstaetten side plates, and intragranular plates. Carbide precipitation converted these morphologies into their bainitic equivalents with increasing isothermal reaction time. Conversion took place more rapidly as the reaction temperature was reduced. The replacement of supersaturated ferrite by bainite, however, did not occur uniformly even within individual austenite grains and was noticeably incomplete at higher transformation temperatures, e.g., 625° to 575°C. The ferritic and bainitic forms of a given morphology thus were frequently found to coexist, often inextricably intermixed, in a given specimen. At 625" and 600°C, the conversion of ferrite to bainite was clearly seen to have occurred entirely by the nucleation of carbides at ferrite/ferrite boundaries, as shown in Fig. 1. The nature of the loci at which carbides formed at lower temperatures was often difficult to ascertain with a light microscope, although the alignment in rows of many of the carbides, illustrated in Fig. 2, suggested that nucleation had again occurred at grain boundaries or subboundaries within the ferrite. Application of the electron microscope confirmed this deduction for bainitic structures formed at these temperatures, as shown in Fig. 3. The dense black appearance of bainite formed at very low reaction temperatures in higher carbon steels3 indicates, however, that isolated lattice defects within ferrite crystals prob-

Jan 1, 1956

By R. Schuhmann

D. W. Fuerstenau (Associate Professor of Metallurgy, Dept. of Mineral Technology, University of California, Berkeley, Calif.) In his excellent paper, Dr. Schuhmann has proposed a mechanism for conlrninution which can be utilized to explain a number of phenomena which have been observed in the size reduction of solids. The difference between the product obtained from a ball mill and from a rod mill can readily be explained by Dr. Schuhmann's hypothesis. Consider that the size distribution of the product from each comminution event is given by y; - 100 (x/ki) (1) where Yi is the cumulative percent finer than size x, ki is the size modulus, and is the distribution modulus for the single comminution event. The total product from a cornminution operation will be the summation of the products from all of the individual cornrninution events. If the size moduli from the various comminution events range from kl to kn , the size distribution of the total product will be a straight line on a log-log plot only to a particle size equal to k ,. Considering 100 volurnes of feed, the value of y for x equal to kl is Riven by where y is the cumulative percent finer than size kj, z is the number of comminution events of kind i, and vi is the volume of each particle crushed. y=z1 qv1+z2 qv2 (k1/k2) +.............zn qvn (k1/kn) (3) F or particle sizes greater than k1 7 e. 6. x = k3, the value of y will be given by k3 A y = z1 qv1 + z2bv2 + Z3uV3 + zn zn qvn (k3/kn)(4) Only when x is equal to k, will y equal 100. 'Thus, the total product from a comminution operation will be a straight line on a log-log plot strictly up to a particle size equal to kl and above that size the line will curve so that y will equal 100 for the coarsest particle or k, (see Fig. 1). The actual shape of the size of distribution curve will depend upon the relative number of conmninution events of different kinds. In the case of ball milling, the ratio kn/kl may be pite large and there will be considerable deviation at the coarse sizes in a Schuhmann plot. kIowever, the grinding action in a rod mill is such that fine particles are protected from the rods by the coarse particles. Consequently, kn/kl will be quite small and the product from a rod mill will give a straight line Schuhmann plot almost to y equal 100%. In his paper, Dr. Schuhmann also has presented a very straightforward definition of grindability derived from the basic energy equation E = Ak-a (5) where E is the energy required to comminute a material to a size modulus of k and A is a constant. According to Dr. Schuhmann, the grindability of a material is the volume of material comminuted finer than unit size per unit of energy expended, this being 1/A. For comparison of grindabilities at any size two numbers are required, both = and 1/A, and anyone tabulating grindabilities should give both of these numbers. Concerning the experimental determination of grindabilities, if the material being comminuted possesses distinct cleavage, the energy re quired for comminution is given by E = Ak-(a-y) (6) where y is small number less than = . Because of cleavage, less energy will be required for comminution. To be certain that cleavage does not affect the gindability of the material being tested, it is suggested that the method used by Dr. Charles for determining grindability be used to ascertain if the deviation exponent y is zero or finite. R. Schuhmann, Jr. (Author's Reply) Dr. Fuerstenau's interesting comparison of rod milling and ball milling

Jan 1, 1961

By Mark Weidenbach

OZ Minerals owns and operates the Prominent Hill copper-gold concentrator in South Australia, built in 2008 and commissioned in early 2009. The concentrator operates via semi autogenous grinding in a thirty four foot ThyssenKrupp Polysius mill followed by a twenty four foot Polysius ball mill in closed circuit with hydro cyclones, designed for an ore throughput of eight million tonnes per annum. The main aspects of designing the circuit are discussed along with the issues faced in the early stages of operation. A review of the optimization strategy applied to the milling circuit such as cyclone optimization, advanced process control strategy implementation and reline management to improve availability is then discussed highlighting areas of significant process improvement that have seen the plant exceed nameplate throughputs above nine million tonnes per annum.

Jan 1, 2012

By Gérald Riverin, Bernard Salmon, Marlo Gagnon, Réemi Asselin, Alain Cazavant, Michel LaBrie

"The Ansil deposit is located in the Noranda mining camp in the southern part of the Abitibi greenstone belt. /twas discovered in 1981 and production began in 1989 with proven reserves of 1.58 Mt grading 7.2% copper, 0.9% zinc, 26.5 glt of silver and 1.6 glt of gold. It is centered at a depth of 1.3 kilometre and represents one of the deepest ore discoveries ever made from surface.The deposit consists of a single massive sulphide lens located at the contact between the Northwest Rhyolite and the Rusty Ridge Andesite. The lens strikes north-south and dips 50° east. The thickness of the lens may reach up to 35 m and can change very abruptly both along strike and along dip. The maximum thicknesses of the sulphide lens are reached in its upper (western) and at its lower (eastern) edges, thus giving the deposit its present shape of two sulphide domes connected by a thinner central portion. The lower dome is underlain by a thick accumulation of massive magnetite which may reach 10 to 15m in thickness. The upper and lower domes are underlain by two distinct pipe-shaped zones of stringer and disseminated sulphides which merge with a larger stringer zone at depth.The Ansi/ deposit displays strong relationships between its morphology and the distribution of metals. Hydrothermal alteration consists of chloritization and sericitization. Chlorite alteration is closely associated with stringer sulphide mineralization while sericite is developed in diffuse zones bordering the chloritic zones. Unlike other similar deposits, the Ansil deposit is characterized by very extensive alteration and sulphide mineralization in the hangingwall of the massive sulphides. Several features of the deposit are consistent with an origin from a submarine hydrothermal system which was particularly long-lived, and which thus promoted the build-up of very high fluid temperatures with extensive replacement (and enrichment) of pre-existing sulphides after their original deposition onto the seafloor."

Jan 1, 1990

By E. D. Gardner

During the summer of 1935 a plan was proposed by which the Government should establish custom sampling plants for gold and silver ores in the mining districts of the West, where the construction of such plants would be justified as an unemployment relief measure. It was known that in many districts prospectors and small operators had small lots of smelting ore that could not be marketed for lack of sufficient capital to get out enough ore at a time to make carload shipments to distant smelters. Moreover, considerable ore of milling grade was available in many districts in which there were no custom milling facilities. It was felt that. accurate sampling and the classification and consolidation of small lots of ore produced from prospects and small mines, with quick cash returns to the shippers on a nonprofit basis, would stimulate the working of small gold and silver properties: Idle mines would be put to work, and there was a possibility that some of the prospects could be developed into regularly producing mines. Each sampling plant was to be centrally located so as to serve the greatest possible number of mines and prospects. Wherever possible, the plants were to be built on a railroad, is ore of smelting grade accumulated, it was to be shipped to a smelter in carload lots. Where a plant was not on a railroad, smelting ore was to be trucked to the railroad after being sampled.

Jan 1, 1936

By Mehdi Razavi

Scope of the Study The possibility and risk associated with mining two stopes at the same time The influence of Creighton Fault on the stability and dilution The effect of irregular shape of stope walls on overall stope stability and ore dilution The support requirements for proposed mining plan

Oct 1, 2010

By R. C. Ku, A. J. McEvily, T. L. Johnston

The microplastic response of a series ofas-quenched Fe-Ni-C martensites has been measured at 77°K. At strains less than JO'3 the flow stress is governed primarily by the transformation-induced dislocation structure of the martensite. Only at strains in excess of 10-3 is the influence of carbon manifested in the flow stress. At these macroscopic strains, typically 10-2, the solid-solution hardening is proportional to (wt pct C)1/3, and, in an alloy containing 0.39 wt pct C, amounts to 50 pct of the flow stress. THE technological significance of high-strength ferrous martensite has stimulated many investigations of its structure and properties. Although our knowledge of the characteristics of martensite has increased immensely, especially with the advent of high-resolution techniques, an understanding of the basic strengthening mechanism still remains elusive. The purpose of the present paper is to consider certain aspects of micro-plastic behavior of Fe-Ni-C martensite which we feel can help to resolve this important problem. Such alloys are particularly suitable for experimental investigation because their compositions can be adjusted to reduce the M, to a temperature low enough essentially to eliminate the diffusion of carbon in the freshly formed martensite.1 The mechanical properties in this condition are of interest inasmuch as they reflect a state that is free of the important but complicating influence of precipitation processes. In this virgin martensite the carbon is distributed as it was inherited from the parent austenite; i.e., it is present interstitially, and gives rise to tetragonality through strain-induced ordering.' In order to determine the source of strength of such alloys, Winchell and Cohen1 investigated the low-temperature macroscopic stress-strain behavior of a series of virgin martensites of increasing carbon content but of common M, temperature (-35°C). They found that the flow stress increased rapidly with carbon content up to 0.4 wt pct; beyond this point the flow stress increased at a much slower rate. It was concluded that martensite is inherently strong. To account quantitatively for the strength of virgin or as- quenched martensite in terms of the role of carbon, Winchell and cohen3 suggested that the carbon atoms, trapped in their original positions by the diffusionless martensite transformation, interfere with dislocation motion according to a model akin to that of Mott and Nabarro. 4 In this treatment, individual carbon atoms are considered to constitute centers of elastic strain and thereby generate an average stress resisting the motion of dislocations throughout the lattice. The additional stress necessary to move dislocations, over and above that necessary for motion in a carbon-free martensite, is given by where L is an effective length of dislocation capable of motion. L was assumed to be limited to the spacing between the twins that are an essential structural element of Fe-Ni-C martensites. They assumtd the spacing to be invariant and of the order of 100A. However, recent work5 has shown that L is variable and can be in excess of 1000Å, so that the assignment of an appropriate value of L is not straightforward. In contrast to the above conclusion that there is an intrinsically high resistance to plastic flow, it has been suggested by Polakowski6 that freshly quenched martensite is in fact "soft" in the sense that dislocations are initially free to move upon application of stress. The high indentation hardness and macroscopic yield stress of ferrous martensites are then a consequence of rapid strain hardening that depends upon carbon in solution. Consistent with this point of view are the results of Beau lieu and Dubé who measured the rate of recovery of internal friction as a function of aging (tempering) temperature in a freshly quenched steel containing 0.90 wt pct C, 0.37 wt pct Mn, 0.1 wt pct Cr, and 0.07 wt pct Ni. The kinetics were clearly consistent with the idea that many dislocations are unpinned in the as-quenched state and that during aging they become progressively pinned by carbon at a rate controlled by carbon diffusion in the body-centered martensite lattice. In order to provide a basis upon which to distinguish between the "hard" and "soft" interpretations indicated above, we have made studies of the initial stages of plastic deformation in Fe-Ni-C martensites similar to those'used by Winchell and Cohen. It will be shown that the results support the contention that dislocation segments in as-quenched material are indeed

Jan 1, 1967

By A. L. Engel

This investigation was undertaken to assess the feasibility of processing an extensive deposit of iron ore near Palisades, Eureka County, Nev., previously penalized for high phosphorus content. The experiments described indicate a satisfactory accomplishment of objective. Chemical treatment will permit a high-phosphorus iron ore to be used. Leaching with acid solution effectively lowers the phosphorus content with a [gligible] loss of iron. The use of hydrochloric acid is preferred because [ ] does not contaminate the ore and the cost is lower. Minus-1/2-inch ore was satisfactorily leached in 24 hours, minus-1/4-inch ore, in 8 hours, and minus-10-mesh ore, in 3 hours. About 40 pounds of acid per ton of ore was consumed in each instance.

Jan 1, 1959

By E. S. Leaver

Experiments dealing with the dissolution of silver by cyanide from various silver minerals were reported several yews ago.4/ This work showed that the silver in some silver minerals is completely and readily soluble in cyanide solution, whereas other silver minerals must be given special treatment before or during cyanidation to obtain much dissolution of the silver by cyanide. Flotation as a method for treating silver ores has not been studied much, except in the case of base-metal ores carrying silver, particularly lead-silver ores. However, many mills now are concentrating gold ores by flotation, and when such ores also contain silver the recovery of the silver by flotation is of considerable importance. Moreover, many mines contain ore in which silver is the principal valuable constituent. Some of these ores are amenable to cyanidation, but the cost of a cyanide plant may be prohibitive because of limited ore bodies or other factors. In such instances the successful application of flotation would mean that the property could be operated.

Jan 1, 1939

By I. Yu. Rasskazov, A. V. Rasskazova, N. M. Litvinova, N. A. Lavrik, V. G. Kryukov

The possibility of the intensification of the gravity processing of complex iron-manganese ore, which contains the finely dispersed inclusions of gold and platinum is shown. Clear mineral fractions of precious metals were sorted out of the different types of ores (quartz-braunite and carbonate- magnetite ore, ferruginous quartzite, carbon-containing shale, breccia-limestone and dolomite breccia). Composition and structures of precious-metals particles was detailed. The industrial content of platinum and gold was found. Precious-metals mineralization is character for all types of ores. High content of platinum is indicated in the carbonate-magnetite (up to 11.3 mg/kg) and quartz –braunite (up to 10 mg/kg) types of ore. Gold is presented in the ferruginous quartzite (up to 1.29 mg/kg) according to the data of the reduced mineralogical analysis. The composition of native platinum depends on the type of the ores: 1) quartz-braunite type contains isoferroplatinum (Pt3Fe) with Cu, Rh, Ir, Ru; 2) carbonate-magnetite type contains isoferroplatinum with Cu, Sn, Ni, Rh, Ir, Os; 3) pyritic-carbon ore contains isoferroplatinum with Sn, Ni; 4) ferruginous quartzite contains isoferroplatinum without admixtures. The ultrasonic treatment of mineral aggregates (P = 41.5 Watt, I = 15 W/cm2) made possible the liberation of fine particles of precious metals. Smooth-shaped or crystalline mineral aggregates are effectively liberated after acoustic treatment. Their surface is freed from impurities. The aggregates with the complex form and structure, which consist of the numerous plates and the grains, are not effectively liberated from impurities. It is shown that the ultrasonic treatment of aggregates with the fine inclusions of precious metals contributes to the liberation of the particles of platinum and gold and intensification of the gravity processing of refractory ore. This approach contributes the increase of the raw-material base of precious metals.

Jan 1, 2018

By B. R. Broekman, K. J. R. Ford

Over the last decade significant development and use has been made of a simulation model that describes the HCFeMn smelting operations in Assmang Ltd. This spreadsheet based, semi-empirical, predictive model is derived from data from the submerged arc furnaces at Assmang?s Cato Ridge, KwaZulu-Natal smelter site. Departures from the empirical base case mass balance due to changes in the ores, reductant feeds and operating parameters are accounted for by use of fundamental relationships and heuristic type ratios which are applied in an iterative manner in the model. The model performs a sequential set of calculations to iteratively converge on a fully consistent mass and energy balance solution, which predicts Mn alloy and slag output masses and compositions for a given set of feeds. 440 MWh per operating day is used as the basis of energy input. The paper describes the key model components, assumptions made and calculation approach. Added to the mass and energy balance is a cost and revenue balance, which allows prediction of the economic performance for each simulation. Substantial use of the model has been made in simulating the performance of the current Assmang HCFeMn furnaces for both smelter investigations and future production planning. The model has also been a valuable tool in the strategic analysis of the use of future Mn ores and has formed the vital link between mine ore grades (metal and slag elements) and the expected alloy output, grade and profitability. Additionally, the model has been used in benchmarking exercises to predict other Mn smelting operations with remarkable accuracy. The paper highlights some aspects of the approaches taken in using the model to benchmark Assmang?s own furnace performance and cost competitiveness with other producers. For Assmang the development and application of this modelling technology has led to transformation in understanding of the Assmang HCFeMn furnace operations and increased awareness of the potential of its Mn ore bodies.

Jan 1, 2004

In attempting to vegetate pnob!em sites some basic pn.i,ne%ptes, namely the phys.ica.e and chemi.ca2 pnope/ti;es og the problem mateni.al and the selection oU hu.itab.2e p!antts {cor pnopagati.on, have to be considered. Pnobtem sites such as tai ing dams, mined-out atteas, and areas which have been genehaUy advehse2y a46eeted by mining activities, are not only o~ten located in a divehs.ity o4 climatic regions, but they also usuaUy dL64en substantially in theAjt .int' Lns c phys.ica2 and chem.i cat pnopen t i.es . To estabP.ish vegetation on such sites 6requent2y requites: .%) The ame Lonaton oU adverse {yaetors in onde t to exeate a hoot env.utonment suitable Uon "satis1actony" plant growth. .%i) These.2eationo4 plant species suited to the improved Hoot env.utonment, and eompat%b&e with the e2imate o{y the site, and the ultimate Land use o4 the site. iii) The estabL shment o{y a Laboratory and p-LLot scale testing programme to aeeomptizh the above.

Jan 1, 1976

By R. I. Klintworth

The multi-reef area of Cooke 3 Shaft provides a unique mining challenge on account of the very thin middlings between the pay bands and the marginal values of the reef. It is uneconomic or impractical, because of the excessive heights, to mine the middlings using conventional stoping methods. The development of suitable mining methods is discussed with particular emphasis on the control of water inflow and the policy of not jeopardizing or sterilizing other reef horizons by the mining method implemented. No definite solution is currently available, but with the aid of instrumentation, computer analyses and mechanized mining equipment, the mine continues to attempt to maximize extraction and profits.

Jan 1, 1988

SOME 70 miles south-east of the of Auckland lies the town of Waihi-the centre of the Waihi goldfield, at present the mining pride of New Zealand. That the glory of most mining camps is evanescent forms an axiom to which it may be expected Waihi will prove no exception. For the last fifteen years, however, Waihi has been an important gold-silver producer, with an annually increasing output, so that now the area ranks as one of the leading goldfields of the world.

Jan 1, 1911

By H. B. Charmbury

In the United States and in Pennsylvania in particular the coal mine drainage problem is of continuing and growing interest to the coal mining industry, the state government, the federal government and the public at large. The 1965 amendments to Pennsylvania's Clean Streams Law prohibit the discharge of acid mine drainage into the streams of the Commonwealth after January 1, 1967. Because of this new legislation, the coal mining industry and state government are extremely aware of the necessity for new techniques for treatment and prevention of the flow of acid mine drainage from the active mines of the Commonwealth in order to keep the coal mining industry in business in Pennsylvania. Approximately one-third of the total volume of mine drainage presently polluting the streams of the Commonwealth originates from "active" coal operations and two-thirds comes from "abandoned" mine discharges. As early as 1963 it became evident that the Commonwealth of Pennsylvania and the coal industry in Pennsylvania would have to take active steps to abate the mine drainage pollution of Pennsylvania's streams and rivers. This became quite clear during the deliberations in Harrisburg which climaxed a twenty year struggle for effective strip mining legislation. The bituminous strip mining law enacted that year made it unlawful to have any acid discharge from active stripping operations.

Jan 1, 1968

By Heidecks C

Sub-level stoping with consolidated rockfill is the main method for mining steep dipping baryte and fluorite veins in Germany. Drifting is fully mechanised using drill jumbos, LHD units and shotcrete support. Stopes have heights up to 50 m with sub-level distances up to 15 m. Winning of the pillars between sub-levels is done by longhole blasting in slices of 3-5 m length on strike. Mucking is done on the lowermost level by LHD units. After each blast rockfill with binder (cement and/or fly ash) is dumped into the scope from the uppermost level. Economic success depends on both geological and technical parameters and factors, eg vein thickness and dip, ore grade, size and distance of sub-level drifts. Of particular influence are ore loss in pillar winning and dilution both in drifting on vein and in pillar mining. A comprehensive computer program has been developed for calculating the total cost per tonne run-of-mine ore respectively per tonne of concentrate for a single stope. The cost model includes development, longhole blasting, mucking and haulage to the mill, milling and ventilation. All costs (labour, energy, supplies, maintenance, depreciation and capital cost) are taken into account. Dilution and recovery are both considered. The program can be used to estimate the cost of mining a particular stope. It can also be used to investigate the influence of relevant parameters and factors on the economic result. By this way it should be possible to optimise the technical design of a stope.

Jan 1, 1995

By W. R. Bruce, R. W. Wittkopp, R. L. Parratt

Introduction The Relief Canyon gold deposit is about 24 km (15 miles) east of Lovelock at the south end of the Humboldt Range in northwestern Nevada. The deposit, is in the Relief-Antelope Springs mining district, which has historically produced silver, antimony, and mercury. There is, however, no mention in the literature of commercial gold production. Fluorite prospects at the gold deposit site have had no reported production. At Relief Canyon, the Late Triassic Grass Valley formation overlies and is in fault contact with the Late Triassic Natchez Pass formation. Epithermal disseminated gold mineralization is found within the various types of fault breccia between these two formations. Geology The Natchez Pass formation of Late Middle to Late Triassic age is composed of more than 300 m (985 ft) of massive gray to dark gray locally carbonaceous dolomitic limestone. Some minor beds of shale and siltstone up to 1 m (3 ft) thick are found in the project area. The limestone is locally silty or sandy. The color of this formation below the oxidation base ranges from gray to black and appears to be a function of carbon content. The Grass Valley formation of Late Triassic age is composed of more than 200 m (655 ft) of interbedded units of thinly parted argillite, hard gray to brown quartzite, siltstone, and shale. Within the oxidation zone, these units are olive gray. A few beds within this formation are slightly calcareous and a number of sections, especially those containing shale, are dolomitic. Below the oxidation zone, the quartzite beds are often slightly carbonaceous and the argillite, siltstone, and shale beds are often highly carbonaceous, giving them a black color. Two types of intrusive rocks have been recognized at the Relief Canyon deposit. Both appear to predate mineralization. Fine to moderately fine grained quartz monzonite dikes, up to 3 m (10 ft) thick, were encountered in several drill holes. In a number of intervals, these dikes have undergone either propylitic or argillic alteration. The age of these types of dikes is not known. It appears, however, that they are either Jurassic or Cretaceous. No gold mineralization has been found in this type of dike. Diabase dikes were also encountered in a number of drill holes. These dikes have almost always been propylitically altered. Although the exact age of the diabase dikes is not known, they are probably equivalent in age to the quartz monzonite dikes. Quaternary alluvium is found forming fans at the base of steep slopes and as recent fill in present day drainages. The alluvium is composed of either Natchez Pass limestone or Grass Valley quartzite and siltstone, depending on which unit served as the bedrock source. A significant portion of the Relief Canyon deposit is covered by Quaternary alluvium. Figure 1 shows a generalized geologic map of the Relief Canyon area. At the deposit's site, the Grass Valley formation appears to have been thrust over the Natchez Pass formation. The age of the thrust is probably correlatable with the Nevadan Orogeny, which gives it a Jurassic-Cretaceous age. The general strike of the thrust, referred to as the Relief Fault, is in a northwest direction. The strike of the bedding of both the Natchez Pass and Grass Valley formations roughly parallel the strike of the Relief Fault. The general dip of both the Natchez Pass and Grass Valley formations is in a southwest direction. The general dip of the Relief Fault, in the area of the Relief Canyon gold deposit, varies and has the appearance of a northeast-southeast striking anticline that plunges in a southwest direction. A small fold perpendicular to the plunge of this anticline forms a dome over the southerly portion of the Relief Canyon deposit. A number of northeast and northwest trending normal faults slightly offset the Relief Fault. Because of their small displacement, they are not shown on the generalized map. Gold Mineralization Gold mineralization occurs along the highly brecciated fault contact between the Natchez Pass and Grass Valley formations. Weak gold mineralization often occurs up to 2 m (6.5 ft) above the thrust in the Grass Valley formation. Most of the ore grade mineralization, however, is present below the Grass

Jan 11, 1984

By Stuart J. Schofield

Introduction The study of the ore deposits of British Columbia is one of fascinating interest not only from a scientific but also from a commercial standpoint and certain wide problems in connection therewith have interested me for some years. Several of these problems have been set forth in papers presented before this Institute on previous occasions. In 1918 it was pointed out that the commercial ore deposits of British Columbia, with one exception, were associated with the granodiorite batholiths of upper Jurassic age. This was followed, in 1922, by a discussion of the ore deposits of British Columbia in which the mineral belts of the province were defined and named, showing a zonal arrangement of ores in which the Pacific coast deposits are characterized by the presence of copper, while the interior belt is marked by the presence of silver, lead, and gold. Methods Of Investigation In reading various reports on the economic geology of British Columbia one is struck with the fact that the mineral veins may be arranged as a system and further examination will show that such a system follows two main directions, one northeasterly and the other northwesterly, both related in origin to the same force.

Jan 1, 1925

By E. Topuz

The objective of mine ventilation is to provide suitable environmental conditions in working places at an economical cost. Maintaining a stable air flow, i.e., a flow quantity and the direction, is of vital importance. In practice, however, air flow quantity in underground openings varies continuously. Therefore, the stability of air flow means relative stability that is the flow in a branch moves along a certain direction while the variation in air quantity stays within acceptable limits. Many underground mines have low stability of mine air flow because almost all types of ventilation systems involve complex diagonal branches through which air flow quantity and, more importantly, the air flow direction may change easily. This creates extremely hazardous, unstable conditions, especially during emergencies. This paper deals with the directional stability of air flows in underground mines in single or multiple fan systems. It examines the rules of directional changes of air flows in diagonal branches, and proposes a method of calculating the stability conditions. It provides a criterion that may be used in designing new mine ventilation systems to ensure a desired degree of stability or in improving the existing systems.

Jan 1, 1986