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By James S. Redpath

"The ways of the world be many and most of the ways be tried, you follow in the footsteps of others till you come to the great divide". These words, written by Robert Service a long time ago, come to mind when I think of our original attitude upon entering into a contract with the International Nickel Company to sink the deepest shaft in the world outside of the Union of South Africa. This shaft was to be sunk to a depth of 7,137 feet from surface, and while it was recognized by all concerned to be a somewhat unusual undertaking in our part of the world, the fact that quite a few shafts had already been sunk to this depth in South Africa probably gave us the feeling that we were following in the footsteps of others. In retrospect, because this shaft was being sunk under entirely different conditions than those existing in South Africa, we were in fact venturing into an entirely new area of endeavour without properly recognizing this fact. It has occurred to me that others considering a similar type of job could benefit from our experience, especially in an assessment of the risks involved and how they compare to those encountered in an ordinary shaft. For comparison purposes, I would define an ordinary shaft as being sunk to a depth of 5,000 ft or less, and a deep shaft as one being sunk to a depth of between 5,000 ft and 8,000 ft. The sinking of the Creighton No. 9 Shaft has been de-

Jan 1, 1997

By T. L. Brekke, T. R. Howard

Stability of an underground opening can be defined in several ways. To geologists and engineers that are used to working in geological environments where the major parts of the works after careful scaling can be left unsupported during the construction period as well as during operation, any rock mass behavior that requires reinforcement, support or lining may be regarded as a stability problem. To colleagues excavating underground openings in geological environments that require continuous use of reinforcement or support and lining "stability problems" are often associated only with those parts of the work where "standard" construction procedure and method of reinforcement or support and lining is inadequate, and special measures therefore needed. In either case, a major objective is to assess the stability problems correctly -as far as possible at the planning stage -- and choose methods for excavation and reinforcement or support that will bring the opening safely and as economical as possible through the problem areas. INTRODUCTION It is not uncommon that stability problems associated with faults or seams have been described as "unexpected." This term has unfortunately been used to indicate that no one knew that the fault or seam existed prior to excavation. A careful review of documented cases where serious stability problems have occurred revealed that the existence of these zones most often was known, but that their behavior was incorrectly assessed prior to or during construction. Miscalculations have been made in terms of the type and magnitude of loads that developed, and in particular in terms of the time-dependant behavior.

Jan 1, 1997

By R. H. Grau

This Bureau of Mines report describes methane liberation trends from all U.S. coal mines liberating over 100 Mcfd (100,000 cfd) (2.8 Mm3/d). Detailed are results from the latest Bureau survey performed in 1985 and shown are mine and company name, location, coalbed and methane liberation quantity. The report shows that in 1985, there were 180 mines producing more than 100 Mcfd (2.8 Mm3/d) and the total quantity of methane liberated by these mines was 303,900 Mcfd (8,605 Mm3/d). This data is compared with previous studies performed in 1971 , 1973, 1975, and 1980. From 1971 to 1985, there was a 10% decrease in the number of mines liberating at least 100 Mcfd (2.8 Mm3/d) but there was a dramatic 34% increase of liberations from these mines. The increase seems to be caused by new mines being developed in deeper, gassier coalbeds, particularly in Alabama. Likewise, in comparing the ten highest liberating mines in 1975, 1980, and 1985, an increase in methane liberations each year was experienced. The report classifies mines by three levels of liberations: greater than 100 Mcfd but less than 500 Mcfd (2.8 to 14 Mm3/d); 500 Mcfd to 1 ,000 Mcfd (1 4 to 28 Mm3/d); and greater than 1,000 Mcfd (>28 Mm3/d). It was found that 87% of liberations were from mines producing over 1,000 Mcfd but these accounted for only 38% of all the mines in the study.

Jan 1, 1987

By G. G. O. O. Uwadiale, M. A. U. Nwoke

The recovery of iron from low-grade (31.6% Fe) Muro iron ore was investigated using reverse anionic flotation of quartz. Tests were performed to investigate the effect of pH, the effects of using distilled vs. tap water as the flotation medium and the effects of using cassava starch vs. dextrin for the depression of iron minerals. The optimum results were obtained around pH 10 and 11, and cassava starch was superior to dextrin in depressing iron oxides. Investigations at pH 11 using cassava starch and distilled water resulted in a concentrate grade of 69.4% Fe with a recovery of 84.4% Fe. Under the same conditions, the use o f dextrin produced a concentrate grade q f 65.9% Fe with a recovery of 78.6% Fe. Investigations at pH 10 using cassava starch and distilled water resulted in a concentrate grade of 60.2% Fe with a recovery of 83.9% Fe, and under these conditions the use of dextrin produced a concentrate grade of 56.1 % Fe with a recovery of 81.5% Fe. Flotation studies conducted at pH 11 using tap water and cassava starch resulted in a concentrate grade of 70.4% Fe with a recovery of 92.7% Fe. This demonstrated the superiority of tap water over distilled water in the reverse anionic flotation process. Investigating the effect of calcium ions on the activation and flotation of quartz with oleic acid revealed that a calcium ion concentration of 50 to 100 ppm was optimum.

Jan 1, 1996

By C. F. Meyer

The recirculation of mine air in an underground colliery, has always been regarded as a hazardous practice in South African collieries. The Mines and Works Act and Regulations states that fresh air, flowing Into an underground section, must flow out directly into the return airways after being utilized once in the section. One of the largest producing underground collieries in South Africa, Initiated a project on controlled recirculation with the permission of the South African Government Mining Engineer. The project involves the recirculation of 65 m3/s of air inside a ventilation district which serves a total of six production areas. The aim of the project was to increase the airflow to the sections without altering the existing ventilation system. Tracer gas studies were carried out to determine the actual distribution of the recirculated air to the individual areas. The results showed that an effective 54,18 m3/s of the recirculated 65 m3/s reaches the sections. This gives an air utilization figure of 83,4 per cent. Dust surveys were also carried out as part of the project to determine the effect of the recirculation of the return air on airborne dust levels. The results showed that the dust levels remained constant in the main intake and return airways and that no hazardous situation was being created. The project proved to be successful and showed that controlled recirculation can be executed and used as a solution for the supply of air to sections without endangering any lives.

Jan 1, 1993

By Jerome F. Machamer

Ninety percent of manganese is used as an alloying element in steel. Smaller amounts are used in dry-cell batteries, in aluminum and bronze alloys, and many other uses. The element has several oxidation states, and oxidation-reduction reactions control its chemical behavior. Manganese, in primary minerals such as rhodochrosite, rhodonite, and braunite, is in a divalent oxidation state, and manganese in secondary minerals, such as pyrolusite and cryptomelane, is in a quadrivalent state. Virtually all manganese ore production comes from three types of sedimentary manganese deposits. First, lenticular beds of manganese carbonate or carbonate-silicate rock in Early Precambrian volcano-sedimentary complexes in South America and Africa have been major sources of manganese ore. Second, laterally extensive beds of mangano-silicate and manganiferous carbonate minerals in certain banded iron formations are major sources of commercial ore and constitute the largest resource of manganese. Finally, manganese accumulations deposited in marginal to epicontinental basins are the largest source of current manganese ore production. Residual deposits resulting from the destructive weathering of rocks with a low manganese content have been, but are no longer, important sources of manganese ore. Most deposits have been altered to some extent by super- gene oxidative processes, and are economic because of that alteration. Outcrops of manganese ores are invariably oxidized and can be very deceiving as to the character of the underlying material. Nearshore facies of epicontinental basins, and nearshore age-equivalents of metalliferous black shales are among the targets for new manganese exploration. Very high-grade manganese occurs in sedimentary deposits altered by metamorphic and metasomatic processes.

Jan 1, 2001

By John H. Duke

The success of most attempts to evaluate gold orebodies depends largely on the reliability and availability of representative sampling data and the accuracy of geological interpretation. An accurate prediction of the continuity of geology is paramount to a reliable prediction of gold reserves. Many of the problems encountered in gold mine evaluation and planning revolve around the difficulties in interpreting complex geology from sparse drilling data and the eagerness of mine financiers and promoters to start mining in order to derive an early cash flow. Following a program of sampling, geophysics, mapping and geological interpretation, a conceptual orebody model evolves in the minds of those working on the project. It is this model which is used typically in ore reserve computations and mine planning. Computer modelling techniques to represent and evaluate the conceptual model are now available. The main advantages of using computer modelling methods lie in being able to rapidly evaluate, modify and quantify the conceptual orebody model. Ore zones or domains can be recognised and defined both from geology and sampling results. Once these domains are defined as a series of boundaries, they may be modelled as two dimensional grid surfaces or as wire-framed three dimensional geometries, depending on the complexity of the orebody. It is the definition of these domains which provides the sound geological basis for reserves computation. To estimate reserves it is necessary to determine both volume of ore and grade of ore. The ore volume is calculated within each domain as defined by the confining surfaces or wire-frames. The ore

Jan 1, 1987

By Eckart Schneider

PROJECT The Walgautunnel is the heart of a new hydro scheme in Western Austria, which is supposed to go into operation in 1984. With it's length of 21,4 km, a bored diameter of 6,25 m and an internal operating pressure of 13,5 bar it will be the largest water conducting tunnel in Austria. Even compared with some of the country's world-famous highway tunnels like 'Tauern' or 'Arlberg' it has to be regarded as a major tunneling project. Owner is the Vorarlberger Illwerke AG, a power corpo¬ration founded in 1924. In the course of the past 6o years this company developed the 'Werksgruppe Obere Ill - Lünersee', consisting of 6 major power plants and 4 large reservoirs. The installed capacity is 1.114 MW in the generating and 529 MW in the pumping mode. The total power output amounts to 1.916 GWh. The water for power generation comes from the glaciers of the Silvrettagroup, as this corner of the Austrian Alps between Vorarlberg , Tyrol and Switzerland is called. The Walgauwerk utilises all the water accumulated by the existing schemes as well as the so far unused run-off water, passing by in the river Ill. The intake to the Walgau-Tunnel is situated right beside the lower reservoir of the recently completed pump-storage plant 'Rodund II' at an elevation of 64o m above sea level. Though the head is a mere 162 m, 1 billion m3 of water will generate 356 GWh per annum. The power-output of the two Francis turbines is 86 MW.

Jan 1, 1983

By Thomas J. Rudolphi

Efficient transport of bulk materials by belt conveyor systems is an important engineering design issue, especially as conveyors become longer and power requirements increase. It is well known within the industry that the parasitic energy loss of a belt conveyor, due to the indentation of the idlers into the belt backing material can be upwards of half the total energy required to drive it on a horizontal flight. Other loss sources includes the rolling resistance of the idlers, belt/roller misalignments, material trampling, frictional and acceleration losses on loading, etc., but energy absorbed by indentation of the belt backing as it passes over each successive idler is the dominant power consuming factor. INDENTATION LOSS MODELS Models to predict the rolling resistance of a cylinder (idler) on a viscoelastic foundation (backing material) have been available for some time. As a contact problem, it is nonlinear (contact length or indentation depth is a dependent on the load and vice versa) in load/deformation response. Rubber backing material is generally assumed to behave as a linear viscoelastic solid, even though it is known that most rubber compounds retain linearity only at very low strains. Because completely general solutions are impossible, practical solutions require simplifying assumptions. Those assumptions relate naturally to two separate pans of any analytical model: I. The deformation model-aspects of the problem relating to the geometry, boundary conditions and kinematics of how the material deforms. 2. The material model-aspects of the problem relating to the material properties, or constitutive relationships, i.e., the stress/strain relationships.

Jan 1, 2008

By T. P. Lim, A. J. Vivyurka, N. K. Davé

Chemical and radioisotope stratification studies were carried out in an abandoned uranium tailings pile at Elliot Lake, Ontario, Canada. The tailings, with an average thickness of 7 m over an area of 16 hectares, were deposited approximately 20 years ago on a layer of black peaty material which overlays deposits of permeable glaciofluvial sands. Geohydro-chemical investigations of this tailings area have been ongoing since 1978. The radioisotope profile study was conducted on three adjacent sites. Solid core samples, collected with a split spoon sampler at various depths, were analyzed for various chemical and radioisotope constituents. The results indicated by 2 distinct zones of stratification; an upper zone, where the constituents were uniformly distributed, and a lower zone, near the peat-tailing interface, which showed increasing concentration with depth. Within the peat layer, the concentration decreased rapidly with no significant levels of contaminant penetration below the peat layers observed, though the samples were limited. The pH profile showed that the top 5 m of the tailings have been oxidized, the average soil pH in this zone was 3.4 compared to 5 .8 underneath it. The groundwater hydrology of the system was observed to be complex showing downward and lateral flow in the dry season and lateral and upward flow during the spring and periods of heavy rainfall. The results are discussed in terms of a model based on the combined segregation and settling process of the tailings fine fraction, and the geohydro-chemical interaction of the solid and liquid phases.

Jan 1, 1981

By S. R. Wardwell, M. J. Robison

INTRODUCTION The first phase of the Los Angeles Metro Rail Project consists of 7.1 km of subway and stations running through downtown Los Angeles from Union Station to MacArthur Park. Contract A146, Tunnels from 5th/HILL Station to 7th/Flower Station, involves tunnel construction through the heart of downtown Los Angeles. The alignment of the tunnels passes beneath twelve buildings, ranging from a 3-story below grade parking structure to numerous 12 to 14 story buildings. The soil along the alignment is alluvium, consisting of silts and clays, clean sands and gravels located above the ground water table. Because of unstable ground conditions, which led to excessive ground losses in several instances, and because of the potential for excessive building settlements, it became necessary to design and perform a program of chemical grouting. The chemical grouting program was designed such that tunneling could proceed on a one shift per day basis, with chemical grounting being performed on the remaining two shifts. Each day 21,000 to 31,000 liters of sodium silicate grout were injected through sleeveport grout pipes installed 18 m deep through the tunnel face. Tunnel advance rates of 12 m per day, 60 m per week, under full chemical grout cover were achieved. The chemical grouting program virtually eliminated both ground loss and settlement and permitted successful excavation of the tunnels. The purpose of this paper is to describe the chemical grouting program used to control ground losses and settlement on Los Angeles Metro Rail Contract A146. The paper begins with a description of

Jan 1, 1991

By C. B. Manula

This paper presents the results of an application of the PSU/USBM Open Pit Materials Handling Simulator (1) to a large stripping mine in the Southwestern Coalfield. The specific objective of the study was to demonstrate the application of the simulator to a complex multi-pit operation, with emphasis on simulating short term and long term production plans recommended for the mine by Theodore Barry Associates (2). Mine Description The mine is located in the Southwestern Coalfield, which encompasses parts of the states of Utah, Arizona and New Mexico. The lease area presently covers 121500 x 103 m2 (30000 acres) in a strip 37.007 x 103 m (23 miles) from north to south and up to 1609 m (1 mile) in width, located at an elevation above 1525 m (5000 ft). Three gently dipping coal seams are present, of which the upper seam (#3) averages 0-5.5 m (0-18 ft); the intermediate seam (#2) averages 0-1.2 m (0-4 ft); and the lower seam (#1) averages 0-2.4 m (0-8 ft). Overburden is mixed sand, gravel, calcareous clay, caliche and sandstone, with calcareous shale and sandstone burden between seams. Where the seams are covered by less than 6.1 m (20 ft) of overburden, they are generally weathered and oxidized and are not extracted. Stripping ratios have ranged from 3.5 to 6.0 with an average of 5.4. All future plans were based on a stripping ratio of 5.3.

Jan 1, 1979

By Stephen G. Gemell

Gold mining has undergone a massive revival in Australia within the past decade. With very few exceptions, these operations are involved with the development of deposits which have been previously and partially worked by underground methods. In most cases, current operations have involved the open cut mining of near-surface remnants or low-grade hanging wall and footwall mineralisation. A growing number of mines are, however, progressing to the development of underground workings either as extensions of existing open cuts or as "stand alone" projects. These operations vary in scale from the Kalgoorlie operations of North Kalgurli Mines Limited and Kalgoorlie Mining Associates to the small scale "gouging" operations of individuals and partnerships. The individual production rates rarely exceed 200 000 tonnes per year, primarily because of the magnitude of reserves or the style of access available for rehabilitation. The narrow-vein mines were initially re-opened using the traditional methods, in particular shrinkage stoping and cut-and-fill stoping. As a result of experience gained in various ground conditions, certain variants have become popular. This paper investigates the fate of the traditional methods in Australian practice, and reviews some of the recently applied variants. This involves not only a perusal of current practices, but also an assessment of the influence of recent technological development on the economics of narrow-vein mining. Some consideration is also given to the effects of previously employed stoping techniques on efforts to access remnant blocks in the re-development of old mines.

Jan 1, 1989

By Glenn Melchert, Paul Eger, Jon Wagner, Dave Antonson, John Folman

Current approaches to reclaiming acid generating tailings include permanent water cover or encapsulation with various synthetic liners and/or clay. Although these methods can be effective, some ongoing maintenance will be required in perpetuity to ensure that water levels are adequately maintained or that the integrity of any capping system is protected. The successful creation of wetlands in tailings basins offers the possibility of creating a stable environment for the tailings with minimal maintenance. Acid-generating tailings from a massive zinc sulfide deposit in Winston Lake, Ontario, were placed in small cylindrical tanks to examine the feasibility and effectiveness of creating wetlands and mitigating acid and metal release. Two uncovered controls were established in addition to the five treatments, which were done in duplicate. Each wetland tank was 117 cm in diameter and contained 6 1 cm of tailings. Treatments included 61 cm cover of wetland soil, 61 cm of glacial till, 61 cm of tailings, and two tanks with 71 cm of water. One of the water covers included the aquatic macrophytes: Elodea canadensis, Potamegeton sp., and Ceratophyllum demersum. Cattails (Typha latifolia) were planted in the plots with substrate cover, and the initial water level was set at ten centimeters. The pH within the tailings in the treated tanks generally ranged between 6.0 and 6.5 and zinc concentration ranged from 0.02 to 0.1 mg/L. The drainage from the controls had an average pH of 3.3, and 7,700 mg/L of zinc.

Jan 1, 2000

By E. Lefeuvre

Precambien in age, the phospate deposit under study which outcrops to the south forms a synclinal consisting of only one mineralized horizon. The synclinal strikes SSW/NNE, dips from 30 to 60 NNE and is about 5 km wide. The deposit has been explored by : - trenches and pits about 50 m apart in the outcropping area ; - core drill holes forming lines perpendicular to the outcrop. These lines are about 100 m apart. On each line, the drill holes are about 50 m apart The phosphate layer is clearly visible and has then been defined from the geologicia1 description of drill hole logs. The layer has been divided into two parts : a very rich, altered part at the hanging wall and a poor part displaying opposite physical characteristics, i.e. a fresh and hard rock. With open pit mining in view, the deposit has been divided into a 3-dimension network of blocks, each block measuring 50 X 50 X 100 m. The tonnages of rich and poor ore, PO25, MgO, Si02, CaO and R2O3 grades and the tonnages of barren waste at the hanging wall and the footwall have been calculated for each block. Given the dip of the phosphate layer ranging from 30 to 60° and its relative thickness (between 10 and 15 m), it was impossible or inappropriate to carry out direct calculation by 3-dimension kriging : too many neighboring blocks to obtain sufficient date and impossibility to draw up the morphology of the layer.

Jan 1, 1986

By J. W. Stevenson

The need for a separate ventilation department within the management team was recognized by the management of U. S. Pipe and Foundry before the company was purchased by Jim Walter Corporation in January, 1970. Because the coal beds of the deep underground coal properties owned by Jim Walter Resources lay 1300 to 2200 feet below the surface, efficient mine ventilation practices are essential through shaft openings varying from 16 to 22 feet in diameter. The extremely high methane gas quantity indicated to be in the coal bed because of the depth of cover and documented by various U. S. Bureau of Mines tests and demonstrated by early mining proved the need to implement this early management decision. At U. S. Pipe and Foundry mine ventilation was a function of the Mining Engineering Department with ventilation personnel assigned to the mine receiving their day to day work assignments from the mine manager or mine foreman. At Jim Walter Resources, the General Manager - Ventilation reports to the Senior Vice President, Mining Engineering, who reports directly to the President, Mining Division, Figure 1 (Company Organization Chart). Position descriptions are static; however positions are dynamic and consequently the degree of emphasis of specific duties and responsibilities change depending not only upon the aptitude of the individual performing the job but also upon his ability to meet the needs of his supervisors; and ultimately upon his ability to react to the economic forces directing our society. The following describes the duties and responsibilities of the General Manager - Ventilation, Figure 2.

Jan 1, 1980

By V. Enns, E. Soprovich, J. H. G. Laflamme

A mineralogical and geochemical characterization study was carried out on arsenic-bearing Ketza mine tailings to complement a column extraction test program with the purpose of assessing the potential of arsenic mobilization from the tailings. Two types of tailings samples (i.e., exposed tailings from the impoundment area and water covered tailings samples) were subjected to column tests under standing water cover and with no water cover but weekly flow-through flushing. The tailings have a uniform size distribution. The samples are characterized by high Fe2O3 (44.0 to 68.2 %) and As (3.7 to 4.4w%) concentrations. Ferric hydroxides are the dominant species that are mostly amorphous or cryptocrystalline. Arsenopyrite, scorodite, an Fe-Ca arsenate hydrate phase, an Fe-Bi-As phase and ferric hydroxides are the arsenic-bearing minerals. Arsenic is essentially contained in Ca-As goethite, Fe-Ca arsenate hydrate phase and scorodite. Goethite, although containing low levels of arsenic, could be a significant arsenic-bearing phase because of its high abundance. There does not appear to be any significant chemical or mineralogical difference between the original exposed tailings and the material left underwater with no flow-through for over a year. There are differences in terms of CaO, Fe2O3, SiO,, As, goethite, scorodite, calcite and dolomite concentrations between the exposed tailings and the column material after undergoing approximately two-year flow-through and aggressive leach tests. These differences are likely to be resulting from a combination of different column test conditions and sampling or heterogeneity of the material.

Jan 1, 1998

By Jerry Motz, Gary Beerkircher

Through a joint development program between Bougainville Copper Ltd. and Nordberg, the MP1000 cone crusher was developed and tested during the period between 1982 and 1987. Following the successful prototype work, Bougainville Copper purchased an additional three units. A second prototype was installed at Codelco's El Teniente three-stage crushing installation in 1987. This unit has also completed its test period successfully. Three additional units have been sold, making a total of eight, to date. Two will be installed as secondary crushers in two separate large aggregate plants. The first unit installed in the United States was placed in operation at Vulcan Materials Company's plant in Manassas, Virginia in early May 1990. Another unit is being installed in Asarco's Ray Mines expansion. The Asarco installation will be a SAG MILL pebble crusher producing approximately 600 STPH of nominal 1/2" product. Application Objective The objective was to design a crusher that would fit on an existing 7' foundation with minimum modification and yet be capable of twice the production of a 7' cone crusher. In meeting this objective, the power rating was increased to 1000 HP, the maximum design crushing force was increased and the crusher was designed to operate at higher speeds than an equivalent size Symons cone crusher. The MP1000 was also designed to operate with a head throw that can be greater than the equivalent Symons design. Design The crusher design criteria involved the recognition of the crushing force increase required for the finer crushed products than

Jan 1, 1992

By S. K. Kawatra, K. J. Shoop

An analysis of an iron-ore processing plant showed that its performance was affected by seasonal temperature variations, resulting in reduced grinding efficiency during the winter months. Laboratory and pilot plant studies were conducted to determine the cause of the observed loss of efficiency at lower temperatures. Two types of experiments were performed: drop tests and wet grinding tests. Drop testing was performed to determine the effect of brittleness and frozen pore water on the breakage characteristics of the ore. The wet grinding tests were divided into laboratory tumble tests and pilot-plant tests. The combined results of drop testing and wet grinding were used to make predictions on the performance of plant-scale autogenous grinding. From the drop tests, it was determined that dry iron ore, basalt and sandstone did not change in brittleness over the temperature range studied (-25º to 25ºC). However, when the rocks were water-saturated, the highly porous sandstone became considerably more fracture-resistant upon freezing, due to "grouting" of the pores and preexisting cracks by ice. This effect was not seen for either the iron ore or the basalt, because they were considerably less porous than the sandstone. Tumble tests on both laboratory and pilot scales showed that both the addition of fines to the test charge and a reduction in temperature caused a decrease in the amount of new fines produced. This reduced fines production was the result of increased slurry viscosity when either the fines content increased or the temperature decreased.

Jan 1, 1998

By Brian W. Hester

Ore was first discovered at the Kidd Creek mine site during the first few days of December 1963 when a diamond drill hole there cut sulfide mineralization. This event marked the end of the search started 11 years earlier with the conception of a geological idea. Economic forces provided the stimulus needed for the application of this idea but other-events, and legal considerations, also played arole in delaying the ultimate discovery. The location of Kidd Creek is shown in Fig. 1. Management of the then Texas Gulf Sulphur Co. had become concerned in 1951 at the company's declining reserves of native sulfur in Texas. This problem was to be alleviated by new discoveries and development of methods to recover sulfur from sour gas. At the time, however, the solution appeared to lie in the recovery of sulfur from pyritic ores. Canada was chosen as the best country in which to find these ores. Walter Holyk was hired in 1951 to conduct the search. By 1952, he was working in the area near Sherbrooke, Quebec, known as the "Eastern Townships" where mineralization consisting of base metal sulfides in volcanic rocks has long been known. His observations here led him to the then revolutionary conclusion that such mineralization could have formed as part of the volcanic process and not, as was generally supposed at the time, by replacement. On learning of the discovery of similar mineralization near Bathurst, NB, he moved there in the spring of 1953. He was able to utilize his new concept and found the Half-Mile Lake sulfide deposit.

Jan 1, 1991