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By Larry Murphy, Mohammad R. Jafari, David M. Giel, Sharat K. Kalluri

"Built in 1940’s to provide a key roadway arterial between New York and Boston, the Heroes Tunnel is the only highway rock tunnel in the state of Connecticut and is part of the Wilbur Cross Parkway. This 365.0 m long tunnel consists of two barrels, each with two lanes of traffic, and carries over 71,000 vehicles per day. For nearly 70 years, the tunnel has seen millions of vehicles pass through but time has taken its toll and now the tunnel lining, ventilation system and radial section joints are significantly deteriorated and in need of rehabilitation. Owned and maintained by the Connecticut Department of Transportation (CTDOT), CTDOT initiated a unique managed approach to mitigate the risk, social impacts and financial implications associated with rehabilitation of this tunnel. This new approach consists of performing field inspection and conducting a comprehensive study to explore six rehabilitation options, including construction of new tunnels and enlarging the existing tunnels. The main criteria for this study are: Impact on traffic and the resulting delay cost; Construction cost; Construction duration; and Construction complexity. This paper will present the results of the alternative study. PROJECT BACKGROUND Location and Importance Heroes Tunnel is located in the State of Connecticut within the Town of Woodbridge and the City of New Haven near the Town of Hamden border. Heroes Tunnel is the only true highway tunnel in the State of Connecticut. The tunnel is part of the Wilbur Cross Parkway (Route 15), a scenic connection between Hartford and New York. Originally constructed and opened to traffic in 1949, the tunneled portion of the parkway passes through the West Rock Ridge. The tunnel was originally named West Rock Tunnel, but was re-designated as Heroes Tunnel in 2003. Route 15 is a vital roadway within the State transportation infrastructure. In the vicinity of Heroes Tunnel, Route 15 carries approximately 71,000 vehicles daily. Locally, Route 15 provides access to residences, shopping centers, and points of interest, as well as other state roadways. Regionally, Route 15 connects New York to the Hartford area and serves as an alternate expressway route to I-95 and I-91. The tunnel is approximately 365.0 m long and passes through the West Rock Ridge. West Rock Ridge is an 11.3 km long rock ridge located on the west side of New Haven. A network of suburban streets abuts the mountain on all sides and a seasonal paved park road climbs to the crest of the ridge. Figure 1 depicts the tunnel topography and location."

Jan 1, 2016

By Keith Kern, Joseph P. Matoney, Kenneth E. Harrison

INTRODUCTION "When there is an economic need for cleaning fine coal to low ash, the technical difficulties will be solved. The utilization of fines below 28 mesh will determine the degree to which beneficiation will be carried. Whenever these "fines" will be processed into smokeless fuels or used as new materials for the chemical industry or for hydrogenation, a higher degree of cleaning will be economically justified”.(l) This quote from the 1943 edition of Coal Preparation correctly anticipates the use of fines as "smokeless" fuel or as a future component of boiler fuel and a1 so as a raw material for the chemical industry or possible new uses such as for internal combustion engines. Fines utilization depends upon an economic need and possibly neglects to say an environmental need or a government funded development program as the prime mover for development. In any event, this conference is faced with the "need" to beneficiate fine coal and undoubtedly among the processes that will be discussed are the unit processes of future coal preparation facilities. HISTORY OF DEVELOPMENT It is interesting to look at where we have been, where we are now and where we possibly will be in the future. In order to put some limitations on this presentation, we will for the most part confine our remarks to this century and particularly developments since World War II. We also will limit our remarks to the preparation of bituminous coals in the United States, and for the most part, to wet processes. In the 1943 edition of Coal Preparation, clues to the processes being used to beneficiate coal are contained in the Chapter Titles. An extensive treatment of "Hand Picking" was contained in Chapter 8. Chapter 7 contained a description of methods for "Preparation at the Face". Chapter 18 dealt with "Slurries, Sludges, Slimes and Water Treatment" and defined this size of coal as being the size at which "Jigs and Chance Cones are usually dispensed with and other methods such as the hydrator, tables, and froth flotation are employed. Generally the limit is 1.7mm to 3.4mm. In the days before World War I1 the then modern bituminous coal preparation plant contained jigs or Chance Cones for the coarse material +6.35mm, concentrating tables, launders or some classifier type of coal washer for the fines (6.35 x 0.15mm), and froth flotation as the only method of cleaning minus 0.15mm coal which was used very sparingly.

Jan 1, 1988

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

The purpose of this section is to demonstrate the development of a method that we believe can be used for simulating and predicting mine water inflow. It will also show how important it is for personnel who model mine water inflow to have considerable experience with this subject, and with the conversion of hydrogeologic frameworks to conceptual models. Several of the models described previously are said to be three dimensional. However, closer examination reveals that there is considerable limitation in the variation possible in the third dimension. A true three-dimensional analysis would require a very large program. Therefore, it seems more reasonable to develop a procedure by which information from a two-dimensional model can be used for analysis of a three-dimensional problem. For a vertical mine shaft this is relatively easy because most two-dimensional models have an option for axisymmetric flow. If it can be assumed that axisymmetry exists about the centerline of the shaft it can be modeled easily as has been illustrated previously. For a horizontal drift or tunnel the analysis is not nearly as straightforward. However, the flow can be divided into two parts as indicated in Fig. 59. Part 1 considers the flow into the side of the drift while part 2 considers the flow into the end of the drift. To calculate the flow rate into the entire drift from the unsteady flow into a unit length the instantaneous flow per unit length may be integrated over the total length of the shaft. This procedure is shown in Fig. 60. The radial flow through any incremental length of the shaft is q. The total flow rate, Q, into the drift at any particular time is given by: The discharge q is a function of the time during which a particular segment has been draining that is dependent on the location x. In other words, a location at x = 0 has been draining during the entire time of expansion while a segment at x = L has only started to drain. Therefore, the time of draining at any point, x, is given by: [ ] where t' is the time over which the mine has been expanded. To determine the relationship between q and t it is necessary to refer to Polubarinova- Kochina (1962, pp. 566-569) and the equation given for the unsteady flow problem. This is: [ ] where q is the steady flow, q,, q, are the coefficients, and t is the time since drainage began. The expressions for q0, q1, and q2 are given as:

Jan 1, 1986

By Robert L. Sundeen, Richard V. Wenberg

Overshot loaders are loading machines which fill a front-mounted bucket by crowding and bucket lifting, passing the loaded bucket over the machine and throw¬ing the load into a haulage unit to the rear. The load is retained in the bucket as the load mass is accelerated in its path over the machine. At the point of discharge, the bucket is stopped suddenly and the load leaves the bucket, following the throw trajectory past the rear of the loader. Fig. 1 shows a typical overshot loader of this type. Overshot loaders were developed for loading in nar¬row drift headings and drawpoints, where it is impossible to place or pass a haulage unit alongside the loader. Haulage access is to the rear of all overshot loaders. The first overshot loader was developed in the early 1930s. The early machines were rail-mounted, air¬powered, and had small loading capacities. To this day, the same basic design configuration exists, and all suc¬cessful overshot loaders still use compressed air. Bucket sizes range from 0.14 to 0.59 m3 (5 to 21 cu ft). The most commonly used machines are rail-mounted, al¬though crawler versions are popular for special applica¬tions, particularly shaft sinking. Rubber-tired machines have found some application where a greater travel distance is required between the loading and discharge points. CONDITIONS INFLUENCING APPLICATION Roadway Type Rail-mounted loaders are used where soft bottom requires rail support or where the long haul is ac¬complished with rail. This situation usually occurs in civil tunnels or in mine drifts that will continue to utilize rail-bound main haulage after development work is completed. Crawler and rubber-tire mounted loaders are used where the roadway surface is rough and lateral travel beyond the range of rail-bound machines is required. Work Area Dimensions The largest rail-bound loaders can clean roadways as wide as 4 m (13 ft) from a single track. Wider headings require multiple trackage. Headroom require¬ments are dependent upon the path described by the bucket as it passes over the machine. The largest over¬shot loader requires approximately 3.4 m (11 ft) of headroom. It follows that the largest heading that can be driven by overshot loaders from a single set of rails is approximately 4 m (13 ft) wide and a minimum of 3.5 m (11.5 ft) high above the rail. The smallest rail-bound loaders can work efficiently in headings as small as 1.8 m (6 ft) wide by 1.8 m (6 ft) high above the rail.

Jan 1, 1982

By John W. Rozelle, David S. Bolin

INTRODUCTION During the normal course of a mining project, a feasibility study will be prepared that addresses all the aspects of the project including geology and mineral inventories, mining and geo- technical, metallurgical processing, infrastructure, environmental considerations and economic analyses. These studies are prepared for various reasons. Occasionally, a feasibility study will be prepared solely for use within a company as justification for development expenditures, however, these documents are frequently submitted outside the company either to raise funds from financial institutions or from public financing on one of the various securities exchanges. Mineral industry consultants are often called upon to participate in the preparation of feasibility studies or to conduct an independent audit and review of an internally prepared study. As the recoverable mineral resources form the foundation upon which the balance of a feasibility study is based, including the economic analysis of the project, the resource aspect of the study is examined very carefully by financial and securities organizations. The balance of this discussion is directed toward the various geological and resource estimation aspects of feasibility studies that experience has shown to be of particular importance. Much of this discussion will be directed toward open pit gold operations, although the principles are applicable to any type of deposit. 2.0 GEOLOGIC RESPONSIBILITIES IN FEASIBILITY STUDIES The principal geological effort in the preparation of a feasibility study is generally dedicated to the following areas: - Preparation of Geologic Reserve/Resource Estimates - Condemnation of Sites for Facilities, Access Roads, Leach Pads, Waste Dumps and Tailings Disposal - Collection of Geotechnical Data - Provision of Mineralogical and Petrological Data Pertinent to Metallurgical Processing - Selection of Appropriate Bulk Sampling Sites for Collection of Metallurgical Samples Each of these items will be addressed in detail in the following sections. 3.0 RESERVE/RESOURCE ESTIMATION An ore reserve estimate consists of the determination of a volume, or tonnage, and a grade for the deposit. There are numerous cases in the mining industry of projects that have experienced severe difficulties when operational ore grade, tonnage or metallurgical amenability have proven to be deficient compared to predicted estimates. An error in estimation of tonnage generally leads to more long term problems affecting mine life and production schedules. However, a significant error in the estimation of grade, particularly if mined grade is lower than anticipated, can have an immediate, and often disastrous, effect upon a project. The geologic preparation leading to an ore reserve estimate generally follows a logical sequence familiar to

Jan 1, 1987

By Nora M. Hutt, Kevin A. Morin

The Sobek humidity cell has been in use with minor modification for more than 30 years to determine bulk primary-mineral reaction rates. As a result, it has become a key tool for predicting the bulk rate of sulphide oxidation, the Carbonate Mole Ratio (CMR) of neutralization-potential (NP) consumption to sulphide oxidation, and the rates of metal leaching. Most humidity cells are operated for less than a year, with the resulting rates extrapolated into the future for years or decades. The validity of these extrapolations was tested by examining longer trends from cells that have operated for three to seven years, as found in the International Kinetic Database (IKD). Additionally, large batches of cells for two minesites in the IKD (45 and 46 cells each) were examined to determine if regular, predictable trends were obtained across ranges of geochemistry at a minesite. The long-term cells show that the rates of sulphide oxidation and acid generation, based on sulphate production, have a 50% chance of stabilizing within one year, with the remainder fluctuating significantly throughout their test periods. Despite the fluctuations, the CMR, which is critical for defining site-specific or rock-unit-specific ABA criteria, remains relatively steady in many cases between 1.0 and 2.0 as expected from principles of aqueous geochemistry. However, if the sulphate-production rate falls below 10-30 mg/kg/wk, the CMR value can increase significantly. At and below these low rates, physical factors like flow rates of surface waters and groundwaters will determine the rate of NP consumption, rather than geochemical factors. Also, if weekly rinse pH falls to acidic values, the CMR will decrease below 1 .O, reflecting the exhaustion of fast- neutralizing minerals. The two batches of multiple cells also confirm these trends across arange of geochemistry. The long-term cells show that copper leaching often stabilizes within a narrow range as long as rinse pH remains relatively constant. Also, there is no strong dependence of leaching on the sulphate rate. However, the rate of copper production often increases sharply if rinse pH falls to acidic values. This suggests the measured copper-production rate in some cases is not a true primary rate, but a lower secondary rate after secondary- mineral precipitation within Sobek cells. As with the long-term cells, the batches of multiple cells show that copper production often increases gradually with decreasing rinse pH, with no clear dependence on sulphate production. However, high rates can also occur at near-neutral pH, so secondary-mineral precipitation of copper is not necessarily common.

Jan 1, 2000

By W. R. Reed, G. J. Joy, B. Kendall, A. Bailey

"Testing of the NIOSH-designed roof bolter canopy air curtain (CAC) has gone through many iterations demonstrating successful dust control performance under controlled conditions. J.H. Fletcher & Co. (Fletcher), an original equipment manufacturer of mining equipment, has further developed the CAC concept by incorporating it into the design of its roof bolting machines. A new CAC design has been developed incorporating the results of testing of the NIOSH design and the recent Fletcher design. Observations from smoke testing conducted on this new CAC design in a laboratory setting show promise for improved performance. INTRODUCTION Exposure to respirable coal mine dust can cause coal workers’ pneumoconiosis (CWP) otherwise known as black lung. If silica is present in the dust, the miner may develop silicosis in addition to CWP. CWP and silicosis are occupational respiratory diseases that are ultimately fatal and have no cure. The only method of prevention is through elimination of exposure to respirable coal mine dust and crystalline silica (quartz). The current occupational exposure limit (on an MRE basis) for respirable coal mine dust is 2.0 mg/m3 during each shift that a miner is exposed in the active workings of the mine or in mine facilities [30 CFR 70.100 2015]. Recently enacted legislation changes this standard beginning in August 2016 to 1.5 mg/m3 for a full working shift [Federal Register, 2014]. If the respirable coal mine dust contains more than 5% silica, the applicable respirable dust standard is reduced; calculated as 10 divided by the percent quartz present [30 CFR 70.101] so that the effective exposure limit for respirable quartz is 0.1 mg/m3 for a full working shift. Sources of exposure to respirable coal mine dust for the roof bolting occupations include infrequent maintenance and cleaning of the vacuum dust collection system and operating the bolter downwind (in the return air) of the continuous miner [Goodman and Organiscak, 2003]. Recent studies have demonstrated that vacuum dust collection systems are effective in capturing bolter-generated dust, if maintained and operated correctly and if proper face ventilation techniques are used. However, most exposure to respirable coal mine dust for the roof bolter occurs when the roof bolter machine is operated downwind of the continuous miner [Potts et al., 2011; Colinet et al., 2013]. Because of this higher exposure downwind of the continuous miner, the roof bolter is generally limited to working downwind of the miner only once during a production shift"

Jan 1, 2016

By Selamet G. Ercelebi, Mahmut Yavuz

An inherent aspect of longwall mining is moving the equipment from a mined-out panel to a newly developed panel. Moving times may be reduced by detailed preplanning. A very useful tool for planning is MS Project. Using MS Project, one can efficiently plan, manage and coordinate a project from conception to conclusion. The critical path analysis identifies potential critical activities which may cause excessive time delays. Changes to the tasks on the critical path can have significant impact 0.n the completion of a project. Assigning additional resources to these tasks (either moving labor from other tasks to critical tasks or providing labor overtime work) may reduce the move time. Simultaneous activities are the candidates for moving the labor from an activity to another. Each time a change is made, the critical path should be recalculated. MS Project quickly recalculates schedules and lets you see how changes in one part of the project will affect the overall plan. In this fashion it is possible to minimize the longwall move time.

Jan 1, 2002

By D. D. Murphy

The Bureau of Mines has studied the mechanism of shear-flocculation of silica with stearic acid. At pH 6 and 12, shear-flocculation of silica was not observed with up to 10-4M (5 lb/st) stearic acid. However, in the presence of 0.1 M NaCl at pH 6, shear-flocculation was observed at stearic acid concentrations 10-4 M (5 lb/st). Adsorption of stearic acid on silica was studied using stearic acids spin-labelled with a doxyl group. The presence of the spin-labelling group did not greatly affect the shear-flocculation properties of stearic acid. Adsorption of labelled stearic acid on silica was measured as a function of surfactant concentration with electron paramagnetic resonance (EPR) spectroscopy and gas chromatography (GC). The GC results indicate the stearic acid was predominately adsorbed on the surface at concentrations above 10-4 M. From the EPR spectra, the formation of aggregates of stearic acid was apparent at concentrations of approximately 10-414 (5 lb/st). The results are discussed in terms of a possible mechanism for shear-flocculation.

Jan 1, 1993

By S. K. Kawatra

The fly ash from coal combustion often has a high unburned-carbon content. This can make it unusable as a cement admixture, which is the major fly ash utilization market. As a result, high-carbon fly ashes currently disposed of by landfilling, at sometimes considerable expense. In this paper, the use of froth flotation to remove the unburned carbon from fly ash is described. Three different fly ashes were studied, which had Loss-on-Ignition (LOI) values of 6.79, 6.90, and 11.11 %. It was found that the use of an alkanolamide/oil mixture as a froth-conditioning reagent could greatly reduce the amount of collector needed, by increasing the affinity of the collector for oxidized carbonaceous surfaces. For each of the three fly ashes, froth flotation reduced the LOI of the ash to less than 2% which is well below the ASTM specification of no ' more than 6.0% LOI for Class F fly ash. Increasing carbon content of the ash tended to increase the quantities of reagents needed in flotation.

Jan 1, 1996

By Chris A. Fleming, Daan Seymore

The Golden Jubilee Mine in South Africa is the first mine in the Western World to introduce resin-in­pulp technology for the recovery of gold from a cyanide-leached slurry. Located in the Eastern Transvaal region of the country, the plant processes approximately 375 metric tons• of ore per day, of which over 90 per cent is mined in an open pit adjacent to the metaflurgical plant, while the remaining 5 to 10 per cent is trucl<ed in from two small high-grade deposits, located about 10 kilometers from the plant. The gold grade in the open pit is about 1 git, while the values m the two high-grade deposits range between 5 and 15 git. At the current treatment rate, the proven reserves m the three locations will last about another 10 years. Early testwork indicated that the ore was not well suited to conventional filtration processing or to heap leaching because of its high cfay content, and that carbon-m-pulp (CIP) would probably prove to be the most viable metallurgical route. Tcstwork to establish the feasibility of CIP for this application was carried out at Mintek in 1985. This worl<, which was based on batch-leaching and carbon-loading tests, indicated that CIP was technically feasible, but that both the rate of gold loading and the gold-loading capacity of the carbon were inferior to those obtained with Witwatersrand ores. The poor adsorption characteristics of the carbon were attributed to the abnormally high concentration of natural organic compounds (humates and fulvates) in the ore. Batch tests were therefore also carried out with strong-base resins, since resins are known (1] to be far less susceptible to organic fouling than is activated carbon. The results were very promising, and pointed to fast kinetics and a high loading capacity. (Fig. 1 ). Notwithstanding these findings, it was decided to recover gold with activated carbon at Golden Jubilee Mine because the CIP process was, by then, well established in South Africa and elsewhere in the world, whereas resin-in-pulp (RIP) was unproven technology. The r>lant was commissioned in September 1986 and, for 18 months, operated in a carbon-in-leach (CIL) configuration, i.e. with a total of . six leach tanks, the last four of which contained activated carbon. As could have been predicted from the laboratory results, the performance of the CIL plant was generally poor, low gold loadings on the carbon (less than 1500 git) and high losses of soluble gold (0.2 to 0.3 p.p.m.) bemg characteristic of the operation. Consequently, it

Jan 1, 1990

By Milton F. Searl

Introduction The urgency of further development of our domestic energy resources becomes more apparent every day. The production capability of existing facilities, taken in the aggregate, declines perhaps 10% per year; imported energy is expensive if available at all; conservation helps reduce demand but eventually will reach a point of diminishing return; new energy sources are promising but past experience, as with nuclear power, teaches us lead times are long and success is not assured. To approach the energy future rationally, it is essential to have some idea of the size of the nation&apos;s remaining energy resources, their characteristics, and the degree of uncertainty involved in their estimates. Resources are, of course, only one end of a long chain leading to energy supply to the ultimate consumer. They are an essential element, however, and the only one not under society&apos;s control. Given that the resources and a national will to produce the energy exist, it is possible to convert the resources to supply in, generally, an environmentally reasonable manner. However, the process of developing supply takes time. Government must permit access to the resource, and remunerative prices and further evolution of technlogy are required. This paper does not deal with these essentials, but they must not be forgotten. This discussion is limited to coal, oil, natural gas and gas liquids, uranium, and oil shale. Other energy sources are discussed elsewhere in this issue. Crude Oil Conventional crude oil resources are of greatest concern today, given our dependence on imported oil and its rapidly rising price. Estimates of oil originally in place (OOIP) range from 619 billion bbl to 1,094 billion bbl, differing by almost a factor of two (Table 1). This is surprising because the U.S. is a mature petroleum province with over 120 billion bbl produced and over 400 billion bbl OOIP discovered. Differences in the estimates involve differences in the experience of the various estimators, the interpretation of present geologic information and historical statistics, the methodology, and probably the extent (i.e., water depths) for which offshore resources are included. The range shown for the estimate of a given estimator appears to be primarily due to geologic uncertainty. This uncertainty can be measured by taking the ratio of the low and of the high estimate to the expected estimate for that information source. The low estimates range from 0.85 to 0.87 of their respective values. The high estimates range from 1.18 to 1.27 of the expected values. In all cases the high estimate is further from the expected value than the low,

Jan 1, 1980

By Searl. Milton F.

Introduction The urgency of further development of our domestic energy resources becomes more apparent every day. The production capability of existing facilities, taken in the aggregate, declines perhaps 10% per year; imported energy is expensive if available at all; conservation helps reduce demand but eventually will reach a point of diminishing return; new energy sources are promising but past experience, as with nuclear power, teaches us lead times are long and success is not assured. To approach the energy future rationally, it is essential to have some idea of the size of the nation&apos;s remaining energy resources, their characteristics, and the degree of uncertainty involved in their estimates. Resources are, of course, only one end of a long chain leading to energy supply to the ultimate consumer. They are an essential element, however, and the only one not under society&apos;s control. Given that the resources and a national will to produce the energy exist, it is possible to convert the resources to supply in, generally, an environmentally reasonable manner. However, the process of developing supply takes time. Government must permit access to the resource, and remunerative prices and further evolution of technology are required. This paper does not deal with these essentials, but they must not be forgotten. This discussion is limited to coal, oil, natural gas and gas liquids, uranium, and oil shale. Other energy sources are discussed elsewhere in this issue. Crude Oil Conventional crude oil resources are of greatest concern today, given our dependence on imported oil and its rapidly rising price. Estimates of oil originally in place (OOIP) range from 619 billion bbl to 1,094 billion bbl, differing by almost a factor of two (Table 1). This is surprising because the U.S. is a mature petroleum province with over 120 billion bbl produced and over 400 billion bbl OOIP discovered. Differences in the estimates involve differences in the experience of the various estimators, the interpretation of present geologic information and historical statistics, the methodology, and probably the extent (i.e., water depths) for which offshore resources are included. The range shown for the estimate of a given estimator appears to be primarily due to geologic uncertainty. This uncertainty can be measured by taking the ratio of the low and of the high estimate to the expected estimate for that information source. The low estimates range from 0.85 to 0.87 of their respective values. The high estimates range from 1.18 to 1.27 of the expected values. In all cases the high estimate is further from the expected value than the low,

Jan 9, 1980

By E. R. Hoskins

The Gulf Coast States ranging from South Texas to Alabama contain vast resources of lignite ranging in energy content from about 4000 to 7000 Btu/lb (as-received) with sulfur content ranging from about 0.5 to 2 percent. Currently, the strip mining rate is about 35 million tons per year in Texas and essentially zero in the other Gulf Coast states. Estimates of future mining rates are as high as 310 million tons per year by the year 2000 with 240 being in Texas. In this paper, we examine the life cycle of the surface mineable lignice based on the most recent estimates of surface mine recoverable resources with two scenarios. The first indicates that a 310 million tpy rate could be sustained for only 15 years while the second scenario suggests that the 310 million tpy could be sustained for 25 to 30 years. If lignite resources are required beyond those that are surface mineable, the deep basin lignite will be developed. Five strategies for developing deep basin lignite are discussed.

Jan 1, 1983

By Thomas E. Garnar

Mineral sand products have been produced from beach sand heavy minerals for nearly a century. These minerals have specific gravities of 3.0 or higher and are extremely resistant to weathering. The most common economic heavy minerals are ilmenite, leucoxene, rutile, zircon, stauarolite, monazite, garnet, and the aluminum-silicate minerals. Of these, ilmenite is the most abundant. In some cases a deposit may contain only one or two economic heavy minerals whereas in others five or six may be present. Heavy mineral mining began in the late eighteen hundreds when monazite was mined for its thorium content. This was followed by mining activity on the North Florida coast just south of Jacksonville Beach - where ilmenite and zircon were first separated during the WW I (1,8). Modern heavy mineral mining began in Eastern Australia during the late nineteen-thirties and spread to the United States during World War II. It is the purpose of this paper to review the commercial heavy mineral products, their sources, and their uses.

Jan 1, 1985

By J. M. Scharer, Bruce E. Halbert, Pettit. Carol M., Lysane Bolduc

The ratio of neutralization potential (NP) to acid production (AP) ratio is often employed for predicting acidic discharges from waste rock dumps. The available NP was evaluated in laboratory studies of limestone dissolution in simulated acidic mine water. The available neutralization potential strongly depended on the particle size. Using limestone particles of 114" (6.4 mm) or greater, less than 20% of the total neutralization potential was exhausted at the onset of acid conditions. Apparently, the dissolution of limestone became mass transfer limited. Under mass transfer limitations, the rate of pyrite oxidation may exceed the rate of neutralization by buffering minerals. The implications concerning the controls on limestone dissolution were examined using a geochemical predictive model for sulphide mineral oxidation in waste rock. The results have shown that in case of heterogeneous waste rock piles the NP/AP ratio is a reliable indicator for short-tam predictions only. Kinetic data on the depletion rate of the neutralizing minerals and geochemical modelling suggest that waste rock environments with NP/AP ratios as high as 5.0 may turn acidic in the long term.

Jan 1, 2000

By G. D. Hood, C. E. Jordan

The US Bureau of Mines investigated rapid froth flotation through more efficient bubble-particle collisions that improved the overall flotation kinetics. An in-line static mixer efficiently mixed the bubbles, water and ore, selectively attaching these bubbles to the hydrophobic mineral&apos;s surface. As the mixture entered the flotation chamber, the hydrophobic mineral was already attached to the bubbles and was quickly floated away from the remaining ore pulp. The rapid flotation system was evaluated for beneficiation with a western porphyry copper ore containing 0.67% copper and with a phosphate-bearing amine flotation tailings sample containing 9.5% P2O5 in a quartz matrix. Over 88% of the copper was recovered from the ore in a rougher concentrate containing 5.9% copper at a rate seven times, faster than conventional flotation. Sixty percent of the quartz in the phosphate-bearing sample was floated at a rate four times faster than conventional flotation. The unfloated product retained 80% of the phosphate in a 16.9% P2O5 concentrate.

Jan 1, 1994

By A. M. Marabini, V. Alesse, Belardi. G.

An alternative flowsheet was designed for the flotation treatment of a complex sulfide ore from the Sotiel Mine in Spain. The ore contains 0.71% Cu, 2.24% Pb and 5.63% Zn. Two specially developed synthetic collectors, methyl-mercapto-benzoxazole (MMBO) and 6n-propoxy-mercapto-benzo-thiazole (PMBT), were examined. MMBO exerts a specific action on copper and zinc minerals after treatment with copper sulfate, and PMBT exerts a specific action on Pb minerals. Flotation operating variables (pH, frother type, frother quantity and feed particle size) were examined by means of rougher flotation tests to determine the optimum conditions for the new collectors. Once the optimum conditions were determined, a series of open flotation circuits were run on feeds of various particle sizes (obtained by different types of grinding). The objective was to obtain high-grade Cu and Zn concentrates. The results of the open-circuit tests demonstrated that a Cu concentrate grading from 18% to 23% Cu and a Zn concentrate grading from 40% to 48% Zn can be produced.

Jan 1, 1996

By D. Harris, M. Osterberg

This paper examines the economics of acquiring mineral exploration rights by royalty-rate bidding when the royalty basis does not provide for full cost recovery. Bidding takes place prior to exploration and the documentation of actual tonnage and grade. Consequently, it increases considerably the risk to exploration capital when the fraction of cost recovered is small and the product price exhibits large fluctuations. Price and grade risk are demonstrated on a 360-Mt (400-million-st) porphyry copper deposit by Monte Carlo simulation of discounted-cash-flow analysis. Copper price is modeled as a stochastic price process, and costs are based on US Bureau of Mines cost models. When both grade and price risk are considered, the probability for a negative NPV is 0.54, given a royalty rate of 5% applied to a basis that allows the recovery of costs up to 50% of product price. Investment risk could be considerably reduced by defining the royalty basis to allow greater cost recovery. Although such a basis would require society to share in the risk, society would benefit from greater resource development.

Jan 1, 2000

By F. F. Peng, D. C. Yang, G. Yanief, E. H. Cho

This study deals with carrier flotation of cerium dioxide, a surrogate of plutonium oxide, which is the major contaminant at the Nevada Test Site (NTS). The objective of this study was to determine the feasibility of using carrier flotation for remediation of the site. In this study, a - 200-mesh fraction of pure cerium dioxide was used. Oleic acid was used as the collector, while calcium carbonate was used as the carrier. Mechanical-cell tests were conducted to float the surrogate from a mixture containing uncontaminated NTS soil, both in the absence and presence of the carrier. It was found that the addition of the carrier drastically enhanced the flotation recovery of cerium dioxide. The flotation recovery of the surrogate increased from 34% to 65% by adding only 0.5% calcium carbonate (-150 mesh). The recovery further increased to 95% by adding 10% calcium carbonate. The flotation method utilizing a carrier of calcium carbonate seems to have a high potential for the remediation of the soil contaminated with radionuclides at the NTS.

Jan 1, 1998