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By Philip A. Dirige

This paper presents the variability of responses of paste backfill specimens that have been prepared with varying binding agent and binder combinations, when subjected to uniaxial compressive strength testing. A testing program conducted for an underground operating mine to produce cost efficient paste fill recipes using four different binding agents, in various ratios and combinations, is presented to illustrate the responses. These binding agents are Normal Portland Cement Type-10 (NPC T-10), oxidation or sulphide attack resistant binder (Binder A), replacement binder B (Binder B), and replacement binder C (Binder C). The paste fill recipes were prepared at 2, 3.5 and 5% binder content at varying binder combinations, at fixed size and tested in a Material Testing System (MTS) under constant strain rate conditions following cure times of 7, 14, 28, 56, 90, 112, 180 and 240 days. The results of tests suggested that the strength data clearly shows a trend of increasing strength with increasing binder content but not necessarily cure time. It is noted that all the blends achieved dramatic increases in strength up to 28 days of cure and then either increased or decreased moderately beyond 28 days curing, depending on the binder content and binder combination. The moderate decrease is apparent for the 2% paste fill blends and in some binder combinations of the 3.5 and 5% blends, which indicated lackluster performance beyond 28 days curing. This lackluster strength performance indicated that the binder content and combination may not be sufficient for the paste fill to achieve complete hydration and acquire adequate strength performance beyond 28 days. This study indicated that economic paste fill recipes can be designed with confidence using affordable alternative binders. However, it is noted that these alternate binders cannot be used as stand alone binders but should be used in combination with NPC to facilitate hydration of the paste fill.

May 1, 2006

By C. H. Hitchcock

History In approaching the subject of diamond-drilling practice, a brief history of the origin and development of the diamond drill may be of interest. The use of black diamonds for boring holes in rock was discovered in 1863 by Rudolph Leschot, a French engineer. In 1864, he employed a diamond bit to bore holes for blasting purposes in the construction of the Mont Cenis tunnel between France and Italy. The method was introduced and developed in the United States in the late sixties. The first deep hole was drilled for coal in 1870, near Pottsville, Penn., by M. C. Bullock. This hole was seven hundred and fifty feet deep. It can be inferred that the diamond drill was introduced into Canada in 1871, when a bore-hole, known as Harper's bore-hole, was drilled for coal at Springhill, Cumberland county, Nova Scotia; and at the old Silver Islet mine, near Port Arthur, the diamond drill was in use as early as 1872. Diamond Drills While M. C. Bullock was developing the diamond drill for exploratory work, the Sullivan Machine Company, of Claremont, N.H., was developing the diamond channelers and gadders for quarry work. These quarry machines were very successful until the high price of black diamonds became prohibitive. In 1878, Albert Ball, chief engineer of the Sullivan Machine Company, designed the Sullivan diamond drill so well known today. The first of these drills went to the Lake Superior iron ranges. After 1880, use of the diamond drill expanded rapidly and spread to all parts of the world where exploratory work was being done. In recent years, the type of diamond drill has not changed much, but better material is going into its construction. Steel castings have replaced cast-iron. Ball and roller bearings are now used instead of brasses and babbits. Gun lubrication has eliminated the oil-can. These improvements make a drill that can be operated for long periods without expense for repairs.

Jan 1, 1933

By M. Vallée, J. -P. Charette, D. P. Gold

"The Oka Carbonatite and Alkaline Complex is a Lower Cretaceous intrusive body of the Monteregian petrographic province. Carbonate rocks of various types, and associated alkaline rocks of the ijolite and okaite suites, are distributed in a double ring-dyke structure 4 Yi miles long by 1 1/2 miles across. Alnoite and alrioite breccia pipes intrude all rocks of the Complex and indicate, from inclusions, that a cover of Palaeozoic rocks existed at the time of emplacement.The results of radiometric, magnetic and gravity surveys of the Complex are compiled and discussed both with respect to the regional pattern and, locally, the bed-rock geology. The Complex is characterized by high-intensity anomalies.Exploration activities started during 1953, and mining operations began in 1961 on the property of St. Lawrence Columbium and Metals Corporation, at one time the world's largest producer of niobium concentrates. Because of the boudinaged and fractured nature of many of the rocks and the variability in grade of the ore minerals, the role of the mining geologist is both demanding and exacting. Rock types are varied and their mutual relation-ships ambiguous, the minerals are rare and complex, and close geological control of the ore is needed in the mine and in the mill. Closely spaced diamond drill holes are needed for mining control, and the ore-waste contact is defined mainly by assay results.In the exploration for 'carbonatites' in Eastern Canada, aeromagnetic surveys are judged to be the best exploration tool, followed in decreasing order by the search for circular structures on aerial photographs, regional gravity surveys and radiometric surveys."

Jan 1, 1967

By F. D. Ford, Z. Dai

"Vale’s Ontario Operations re-introduced Cu/Ni separation at Clarabelle Mill in 2006. The mill processes ores from a number of mines in the Sudbury basin. A saleable Cu concentrate is produced at 30-60% Cu recovery from mill feed, which has to meet the quality specification. In particular, Ni content in Cu concentrate must be less than 0.6%, otherwise, penalties will apply and diminish the value of Cu concentrate, in addition to the lost Ni value. Generally, a Ni grade of less than 0.4% is obtained in Cu concentrate. More recently, while processing Cu-rich ore, Cu concentrate contained more than 0.6% Ni which is abnormally high. Detailed mineralogical investigation shows that Cu-rich ore contains more millerite in comparison to typical Sudbury Ni sulphide ore. Millerite is found to be more floatable than pentlandite and is much more difficult to depress using high-lime aeration conditioning, which is the standard practice in Cu/Ni separation to float chalcopyrite while depressing pentlandite. The current operation uses sodium cyanide at 20 g/t to depress pentlandite and pyrrhotite. However, it is found that a high dosage is required to depress millerite. It is also shown that starch and dextrin are promising in millerite depression.INTRODUCTIONPrior to bulk smelting in 1991, Cu/Ni separation was carried out at Vale’s Copper Cliff Mill, Ontario Operation. The typical Cu concentrate grade was around 30% Cu and 1% Ni. This is very high in comparison to today’s value of 0.4% Ni and would not be acceptable for external processing. However, the Cu concentrate was processed internally at the Copper Cliff smelter/refinery, so the Ni contained in Cu concentrate was not lost and no actual penalties applied. During the SO2 abatement initiative, bulk smelting was chosen because of the perceived simplicity and consequently, Cu/Ni separation was eliminated (Agar, 1991) and the Copper Cliff mill was subsequently demolished. In 2005 and after more than one decade of bulk smelting, business changes required the removal of some Cu concentrate from Clarabelle Mill bulk concentrate to provide smelter capacity for Voisey’s Bay concentrate in the Copper Cliff smelter. In essence, this strategy calls for Cu removal at Clarabelle Mill (Lawson, 2006). In this case, Cu concentrate is sold externally and it must meet the quality requirement of less than 0.6% Ni."

Jan 1, 2011

By D. Gendzwill, N. Martin

"In early 1987, the Patience Lake potash mine near Saskatoon, Saskatchewan, was closed due to uncontrollable floodingWater inflows were first encountered during routine mining in 1975 at the south boundary of the mine lease. Underground grouting and construction of bulkheads limited the flo but did not completely stop it. Subsequent seismic surveys showed that the mine had approached the edge of a large collapse structure in which the salt and potash had been partly dissolved and removed by groundwater.In 1986, the small flow into the mine started to grow. Despite renewed grouting, the flood increased and outstripped pump capacity. The grouting program temporarily slowed the inflow but was unable to stop it. Eventually, from seismic and drilling evidence, it became clear that the water source was an extensive system of vertical fractures which would be impossible to seal by drilling and grouting from underground. A new plan to install 14 bulkheads underground would have sealed the water source from the rest of the mine but the construction program could not be completed before rapidly increasing inflow threatened to fill the mine. The underground mine was closed and a new solution mine started in 1988."

Jan 1, 1996

By Jim Wickens, Vince Lofftus, Mark Kravetz

"This paper discusses how Bald Mountain Gold Mine (Ely, NV) dramatically reduced mercury production by treating process solutions with a liquid reagent based on organic sulfide chemistry. Cherokee Chemical Co. Inc. (Vernon, CA) markets this reagent as UNR 81 1A.A summary of the bench tests comparing various reagents is given. The ensuing plant trial is discussed and successful mercury removal results are presented. Data is presented to show that early removal of mercury in the process reduces the amount of mercury collected at retort. Data is also presented to show that use of this reagent does not reduce gold recovery. Plant trial treatment costs are discussed. Finally, the potential of this technology to substantially reduce closure costs as compared to traditional mercury removal options is discussed.INTRODUCTIONBald Mountain Mine produces gold. Mercury is a costly by-product of gold production because no revenue is generated from mercury. Traditionally, mercury leaches out with gold and follows the gold throughout the processing circuit until it is eventually condensed and collected at Retort. Minimizing or eliminating mercury production minimizes worker health and safety risks, minimizes environmental liability, and reduces costs. In the long term, minimizing mercury production will make mine closure easier and less costly.Bald Mountain Mine dramatically reduced mercury production by treating process solutions with liquid reagents called UNR 811 and UNR 811A. Several reagents were bench tested before proceeding with a plant trial. The reagent complexes the mercury to form an insoluble organic sulfide precipitate. In independent tests, the precipitate passed acid mobilization tests and is considered stable."

Jan 1, 2001

By Chris Fleming, John R. Goode, James Brown

"Data will be produced from two uranium feasibility studies that were conducted in 2008, both involving uranium recovery from leached slurry by resin-in-pulp (RIP) with strong base resins. Excellent metallurgical results were produced in RIP pilot plants in both studies, and the loaded resins were eluted with strong sulphuric acid and recycled to RIP in an integrated process. Resin losses by either abrasion or mechanical breakage were immeasurably low.Metallurgical results from these two campaigns will be presented, along with an economic analysis of RIP versus the alternative process of CCD or filtration and solvent extraction.INTRODUCTIONThe recovery of uranium directly from pulp with strong base resins in RIP contactors was practiced commercially in the USA in the 1960s, but the process never established itself as a viable alternative to solid/liquid separation and solvent extraction. This was due mainly to the necessity to do a sand-slime separation partially resulting from the unavailability of coarse resin beads, inadequacies in the design of the RIP contactors and screens to separate resin from pulp, as well as problems with the strength and durability of resins in agitated pulps.There have been significant advances in the last 30 years in each of these areas. All the commercial resin manufacturers are now producing very tough, durable, strong base resins, and are manufacturing them in a large bead size that is better suited to RIP applications. In addition, with the emergence and wide spread implementation of the carbon in pulp process in the gold industry in the 1980s, many innovative techniques for preparing leached slurries and separating carbon from pulp have become firmly established. These techniques have proven to be equally applicable to resin separation from pulps in RIP contactors."

Jan 1, 2010

By G. S. Dobby, J. A. Finch

"I. INTRODUCTIONA significant development in flotation over the past few years has been the increasing industrial use of flotation columns, primarily in Canada. The column differs dramatically from conventional mechanical flotation machines, both in design end operating philosophy, which has been a principal reason for its slow acceptance by the mineral industry. Boutin, Tremblay and Wheeler (1. 2) developed the concept in the early 1960's and since then plant and laboratory columns have been marketed commercially by Column Flotation Company of Canada Ltd. In 1980-81 at Mines Gaspe, Quebec, three columns replaced 13 conventional cleaners in a molybdenum upgrading circuit, with superior results (3). Since then many of the copper producers in British Columbia have installed columns for copper and molybdenum cleaning, including Gibraltar, Lornex, Highmont and Island Copper. There now appears to be widespread interest in flotation columns.The column is particularly attractive for applications involving multiple cleaning stages and can upgrade In a single stage compared with several stages of mechanical cells. This results in simpler, more controllable circuits. Importantly, the column itself is well suited to computer control. A significant problem, however, has been that of translating laboratory data to plant performance. The purposes of this communication are to address the problem of column scale-up and to describe some of the transport mechanisms of a flotation column, leading to a kinetics based model.A schematic of a flotation column is shown in Figure]. Industrial columns are up to 13 m high and 0.3 to 1.8 m in cross section (either square or circular). Three zones can be identified:a collection zone, with feed entering 2 to 3 m below the top of the column and descending against rising bubbles generated at a gas sparger; a washing zone, where a packed bubble bed is generated by a downward flow of wash water; and a conventional froth (2 - 5 cm thick), the sole purpose of which is particle transport to the launder. The tailings withdrawal at the bottom of the column is controlled at a rate slightly greater than the feed flowrate (called a positive bias).An important element is the washwater, added just below the froth zone. The component of the washwater that flows downward (to make up the bias) is sufficient to prevent coalescence of rising gas bubbles. The packed bubble beg so generated is very efficient at suppressing gangue entrainment to the concentrate."

Jan 1, 1985

By S. Banisi, E. Arghavani, A. Akbari, Gh. A. Parsapour, S. Darvish Tafvizi

The new copper processing plant of the Sarcheshmeh copper complex consists of two parallel circuits. After a primary crushing, the ore is sent to a SAG mill and the product is further ground in a ball mill. The overflow of the hydrocyclones is fed to a flotation circuit which contains 8 rougher tank cells (RCS130), 3 cleaner cells (RCS50), 5 scavenger cells (RCS50) and a 4 x 12 m flotation column as recleaner. The circuit was initially designed to process a feed containing 0.8% Cu, but because of the ore type change, the feed grade decreased to 0.6% Cu. This resulted in a reduction of final concentrate grade and recovery from 28 and 85.5% to 24 and 84.4%, respectively. Based on the original design, the copper and silica recovery in the cleaner cells should be 69 and 55%, respectively, but these values increased to 85 and 75% because of a higher retention time. The rather high silica recovery (i.e., 75%) was found to be the main source of the lower final concentrate grade. In order to reduce the retention time of particles in the cleaner cell from 13.7 to 6.9 min, the rougher concentrates of two parallel circuits were fed to only one cleaner-scavenger and regrind circuit. This modification increased the cleaner and final concentrate grade from 15.1 and 24.5% to 17 and 26%, respectively. To prevent a possible reduction in the copper recovery, the froth depth in the scavenger cells was reduced. The overall outcome of the circuit modification was evaluated to be a 10% reduction in the energy consumption without any loss in the overall copper recovery.

Jan 1, 2016

By M. Kinne

Catalyst preparation/ characterization Slightly modified DP (~Schumacher et al. Catal. Lett. 89, 109, 2003) CeO2 HAS 15 from Rhodia calcined in air for~ 2h BET: ~188m2/g, particles size: 5.7nm (TEM, XRD) + HAuCl4 at pH of 6.5-7, (>95% of the Au was deposited) ? Au/CeO2 (2.6wt% Au, ICP-AES) A new conditioning procedure was applied, including heating at low temperature in N2 and H2/N2 mixtures) Patent DE 10205873A1. Very small Au particles <2nm, for CO oxidation catalyst as Au/TiO2, Au/Fe2O3, and Au/Co2O3 Here Au particles size <3.5nm (XRD) Summary ? CO is nearly exclusively adsorbed on the Au particles ? Positive reaction orders of 0.5 are found for CO and H2O ? Ea= 40kJ/mol ? The catalyst is very active and stable ? During WGS reaction huge amounts of formates are formed and cover the surface ? The activity of the catalyst seems to be correlated with the formate species.

Oct 1, 2003

By David Scott, Denis V. Lachance, Daniel Marcotie

"LocationThe Heath Steele mine is located 50 km northwest of Miramichi, New Brunswick.Brief DescriptionHeath Steele is a 2700 t/d, copper-lead-zinc operation which employs conventional crushing, grinding, and differential flotation.HistoryThe Heath Steele mine is located in northeastern New Brunswick approximately 50 km northwest of the city of Miramichi and 65 km southwest of Bathurst. The massive sulphides containing ore values in zinc, lead, copper and silver were discovered in 1953 as a result of a joint venture between the American Metal Company and the International Nickel Company of Canada. The discovery was named after Mr. Heath Steele, manager of exploration for Amco at the time of the discovery.Mill production began in February 1957; however, the mine was forced to shutdown in 1958 due to depressed metal prices and metallurgical problems. It was re-opened in 1962 and milling operations resumed at 725 t/d until 1968 when the milling capacity was boosted to 1450 t/d and then to 2700 t/d in 1969.OreThe main ore zones are massive sulphides with very fine mineral grain. The chalcopyrite, galena and marmatitesphalerite occurrences are finely disseminated.Liberation is accomplished by grinding to 68% to 72% <400 mesh. Average mill feed assays are 0.8% to 1.2% copper, 1.8% to 2.5% lead, and 5.5% to 6.5% zinc, with 60 to 75 g/t silver."

Jan 1, 2000

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

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

Jan 1, 1984

By D. C. Sharpstone

THE Polaris-Taku mine is a new gold property, situated in the far north-west corner of British Columbia. Production was started in November, 1937, at the rate of 150 tons a day, and recently has been increased to 200 tons. The property, which promises to become one of British Columbia&apos;s important gold mines, was developed and brought into production in twenty-nine months after work was started on it as a virtual prospect in June, 1935, by Edward C. Congdon and associates. Inaccessibility, extraordinarily complicated geological problems, and difficult metallurgy retarded development of the property before acquisition by the present interests, an::l, in fact, discouraged two companies that had spent substantial sums upon it from continuing development. This paper is concerned with reviewing the development of the property and describing the principal features of the geology, the mining and metallurgical practices, and the equipment employed. Acknowledgments The writer is indebted to Mr. Edward C. Congdon, President of the Polaris-Taku Mining Company, and to various members of the mine staff for their assistance in assembling the data for this paper. Special thanks are due to Mr. George G. Griswold, Jr., of Butte, Montana, who generously and painstakingly prepared the section dealing with milling and power. Microscopic studies of ores were made by Professors Frank Grout an::l G. M. Schwartz, of the University of Minnesota, and by Mr. Griswold.

Jan 1, 1938

By N. Hiroyoshi, R. D. Corpuz, M. Ito, Hokkaido University, C. B. Tabelin, S. Veerawattananun

"Acid mine drainage (AMD), the very acidic and heavy metal-contaminated leachate generated when pyrite-bearing rocks are exposed to surface oxidizing conditions, is a serious environmental problem encountered after the closure of mines and mineral processing operations. In this study, we investigated the effects of soluble phases and metallic oxides, which are commonly found in strongly weathered pyritic wastes, on pyrite oxidation. Strongly weathered pyrite containing melanterite (FeSO4·7H2O) as the main soluble phase was obtained from Furikusa mine, Japan. The sample was washed using two methods to vary the relative amounts of melanterite in the sample. The effect of metallic oxides on pyrite oxidation was evaluated using ultrapure hematite suspensions. Our results showed that under acidic conditions, the effect of soluble salts on the oxidation of pyrite was stronger than hematite. More thorough washing of the sample resulted in the formation of an arsenic-bearing iron-oxyhydroxysulfate coating that minimized pyrite dissolution. This oxidized layer became relatively thicker and mechanically stronger when hematite was introduced after the more thorough removal of soluble salts. Characterization of this “coating” coupled with geochemical modeling suggest that it is most probably composed of two materials: schwertmannite and scorodite. Based on these results, a new approach to suppress pyrite oxidation by directly using iron-oxyhydroxysulfate as coating material is introduced.INTRODUCTION Acid mine drainage (AMD), the very acidic and heavy metals-contaminated leachate produced in closed/abandoned mine sites (Evangelou, 1995), mineral processing and coal cleaning tailings (Evangelou, 1995), and tunnel construction waste rocks (Tabelin and Igarashi, 2009, Tabelin et al., 2012), is one of the oldest and probably the most serious environmental problem of anthropogenic origin. When released to the environment, it rapidly pollutes river, soil and groundwater systems leading to the destruction of affected ecosystems. AMD is produced when pyrite (FeS2), one of the primary gangue minerals in sulfide ores and coals, is exposed to oxidizing conditions as illustrated by the following reactions:"

Jan 1, 2016

By F. R. Jones, B. W. Taylor, A. R. MacPherson, L. J. Lamb

"The Lake St. Joseph Iron Deposit is in Northwestern Ontario. Exploration to date by Steep Rock Iron Mines Limited has indicated an ore reserve of 683 million long tons. Metallurgical investigations have proved that suitable oxide pellets and/or superconcentrates can be economically produced from this ore. Studies of the Lake St. Joseph iron ore deposits have covered the period 1956-71, and have embodied applications of the improvement in mining and processing technology over this period. Market demand in the early 1960's for high-grade oxide pellets as a replacement for natural iron ores for the blast furnace first drew attention to the potential of the Lake St. Joseph deposit. The current market trend toward utilization of sponge iron to replace steel scrap as feed to electric furnaces and cupolas now suggests that the deposit can better be exploited as a prime source of superconcentrates."

Jan 1, 1972

By W. E. Kuhn

Introduction The work, of which some of the results are described in this paper, was started with the object of investigating the properties of certain binary alloys of titanium made from powders. Our main interest was in the possible use of such alloys in gas turbines. We were attracted to titanium, as have been others, by its high melting point (I,724° ± 11°C.), its low density (4.54), and the likelihood that, if a relatively wide use could be found for the metal, it might become available in commercial quantities at a fairly reasonable price. The readiness with which the metal reacts with hydrogen, oxygen, and nitrogen was recognized from the very beginning. At the same time there seemed a good chance that alloys could be produced which, because of the formation of protective films, would resist the attack of these and other corrosive agents. No specifications were drafted to cover the &apos;quality&apos; of the titanium hydride powder which was used in our work and which was, to judge by the results of a few preliminary experiments, the most useful source of titanium then (1944) available in sufficient amount for our purpose. During the course of our work, the need for such specifications became very clear (see Sections 9 to 12 of this pa per). Even today, lack of knowledge regarding the effects on titanium of impurities, some of which may later be looked upon as essential elements in commercial alloys of titanium, and difficulties in manufacture, make the preparation of specifications for commercial titanium and titanium compounds a doubtful procedure. The papers presented at the 1948 Symposium on Titanium, sponsored by the U.S. Office of Naval Research lend colour to this view. At the time our work was commenced, one had to depend upon the manufacturer to supply what he considered to be his &apos;best&apos; material. It may be some time before we can ask for more than this, the question as to what is his &apos;best&apos; being left to the manufacturer to decide. However, in this connection the results of our study may be of some value and on that account we are constrained to publish them.

Jan 1, 1950

By W. G. Miller

An overall study was conducted to evaluate the properties and processes influencing the rate and magnitude of consolidation for oil sands fine tailings produced using different extraction processes. As part of this overall study, consolidation tests using slurry consolidometers were carried out for caustic and non-caustic fine tailings. The influence of a change in bitumen extraction process (caustic versus non-caustic) on consolidation properties, namely compressibility and hydraulic conductivity, was determined, as was the effect of adding a coagulant (calcium sulphate [CaSO4]) to caustic fine tailings. For fine tailings originating from two different oil sands ores (Ore A and Ore B), results were presented in terms of variation in average void ratio with time (settlement), void ratio with effective stress (compressibility), hydraulic conductivity with void ratio and coefficient of compressibility and hydraulic conductivity with sodium adsorption ratio. Findings indicated that chemical characteristics of fine tailings (water chemistry, degree of dispersion) do not have a significant impact on their compressibility behaviour and have only a small influence at high void ratio (low effective stress). The biggest advantage of non-caustic fine tailings and treating caustic fine tailings with coagulant is an increased initial settlement rate and slightly increased hydraulic conductivity at higher void ratios. Thereafter, compressibility and hydraulic conductivity is governed by effective stress.

Aug 1, 2010

By W. G. Miller, J. D. Scott, D. C. Sego

An overall study was conducted to evaluate the properties and processes influencing the rate and magnitude of consolidation for oil sands fine tailings produced using different extraction processes. As part of this overall study, consolidation tests using slurry consolidometers were carried out for caustic and non-caustic fine tailings. The influence of a change in bitumen extraction process (caustic versus non-caustic) on consolidation properties, namely compressibility and hydraulic conductivity, was determined, as was the effect of adding a coagulant (calcium sulphate [CaSO4]) to caustic fine tailings. For fine tailings originating from two different oil sands ores (Ore A and Ore B), results were presented in terms of variation in average void ratio with time (settlement), void ratio with effective stress (compressibility), hydraulic conductivity with void ratio and coefficient of compressibility and hydraulic conductivity with sodium adsorption ratio. Findings indicated that chemical characteristics of fine tailings (water chemistry, degree of dispersion) do not have a significant impact on their compressibility behaviour and have only a small influence at high void ratio (low effective stress). The biggest advantage of non-caustic fine tailings and treating caustic fine tailings with coagulant is an increased initial settlement rate and slightly increased hydraulic conductivity at higher void ratios. Thereafter, compressibility and hydraulic conductivity is governed by effective stress.

Jan 1, 2010

By Matthew A. Perras

With growing populations, infrastructure projects are utilizing underground space to expand transportation networks, water supply and sewer disposal systems and hydropower facilities. These new projects require large diameter tunnels to meet the growing demands. As the tunnel diameter increases, not all rock mass properties are scalable and precedent experience becomes less reliable. One such non-scaleable property is lamination thickness. Numerical modelling was conducted to determine the influence of lamination thicknesses on a 16 m diameter tunnel. Five different areas of excavation response were found to exist for laminations ranging between 0.16 to 16 metres in thickness. For this tunnel diameter, critical lamination thicknesses of 6 metre and 1 metre were found to exist. Above 6 metres the excavation response was similar to isotropic models. Below 6 metres the stresses and yield are channeled by the lamination boundaries, increasing plastic deformation. Below 1m multiple lamination units become unstable and the displacements increase significantly and the shape and extent of the yield zone size changes dramatically. The yield zone is controlled by bed deflections and bed parallel shear. As the lamination thickness decreases tensile failure first begins in the haunch area and progressively extends above the crown until a self-limiting plastic yield zone shape is reach at lamination thicknesses below 0.4 metres. Conclusions drawn from observations at the Niagara Tunnel Project are those of the authors.

May 1, 2009

By A G. Rogers

This paper discusses the lubrication of surface and underground mining machinery and equipment from the standpoint of machine components. The machine components considered are bearings, cylinders, gears, chains, couplings, fluid drives, hydraulic systems, screws and wire ropes. The best selection of lubricants for these machine components, and for sub-divisions of the components, are outlined. They form general recommendations, for average operating conditions in both surface and under-ground mining operations. The American Petroleum Institute&apos;s service classification of both engine crankcase oils and gear oils is de-scribed and used in recommending products for these applications. The practical aspect of lubrication is kept foremost in the recommendations, which are designed to meet the requirements of a wide variety of equipment with the minimum number of lubricants. THE lubrication of mining machinery and equipment presents a variety of problems peculiar to the industry. The type of mine, the ambient temperature, water, dust, and operating and maintenance practices, as well as the age and condition of the equipment, are variables which have to be considered in making lubricant recommendations.

Jan 1, 1963