Search Documents

By R. Woods

INTRODUCTION Gaudin (1), considered that "the mechanism of mineral collection is the central problem of flotation theory". From their work on adsorption phenomena in flotation for more than fifty years (2), Gaudin and his co-workers have made major contributions to the elucidation of this mechanism, a fact acknowledged by their contemporaries in the field (e.g. (3)). Their investigations in a number of areas, e.g. demonstrating that submonolayer quantities of adsorbed collector species can induce flotation, or identifying the influence of substances which can activate or depress (1), have helped to provide the basis for and the development of a theory to account for the interaction between mineral and collector. It was established in the late fifties (4) that the presence of oxygen is a prerequisite for the flotation of metal sulfides with thiol collectors. This fact rendered inadequate the concept of simple ion adsorption being responsible for rendering these minerals hydrophobic. The requirement for oxygen has been explained (1) in terms of this species being a component in chemical reactions between the sulfide surface and the collector. Gaudin has been instrumental in identifying the products of chemical reactions at sulfide minerals (1.5), a line of work which has been continued particularly by Finkelstein and co-workers in South Africa (6). In presenting a detailed mechanism of the mineral-collector- oxygen interaction, it has often been considered that the mineral surface - must first be oxidized to a metal oxy-sulfur species which

Jan 1, 1976

By W. A. Zingg

Impervious zones in copper waste heaps create a problem in contacting the rock with leaching solutions. These zones are normally associated with compacted layers .formed on the surface of each lift by dump truck traffic and by disintegration of the rock exposed to atmospheric conditions. These layers are usually only a few feet in depth--but, nevertheless, act as barriers to vertical flow causing the leach solution to perch in the heap, Consequently, much of the rock remains untouched and the leach solution breaks out the side of the heap instead of continuing downward to bedrock. One solution to this problem is breaking the hard pans with explosive charges. However, to use explosives means adopting new explosive materials and methods. To use this process successfully requires finding answers to operating conditions inherent to waste heaps. Foremost among these are: (1) type and size of casing that can be used because of limitations brought about by drilling problems, caving and, corrosive conditions; (2) perforating the casing at required levels; (3) an explosive capable of being pumped; (4) obtaining detonation of the, main charges through, the wail of the casing; (5) an explosive with the ability to propagate through the voids of the rubble and one which is not affected by water, acidic solutions, or back pressure conditions from hydrostatic heads; (6) to heave and rupture the compacted zones without chambering or recompressing the waste material.

Jan 1, 1970

By D. K. Jorgensen

The Bullfrog and Montgomery-Shoshone deposits, near Beatty, NV, have a combined drill-indicated reserve of 15.8 Mt (17.4 million st) of ore. Average grade is 3.5 g/t (0.103 oz per st) gold and 8.2 g/t (0.24 oz per st) silver. The deposits occur in Miocene volcanic rocks that underwent large-scale movement within a detachment-faulted terrane. The principal host to gold mineralization is a strongly welded rhyolitic tuff that experienced brittle deformation during plate movement. Open space fractures and breccia were filled with quartz, calcite, and adularia accompanied by lesser pyrite, gold, and acanthite. Gold occurs as micrometer-sized particles within and adjacent to the low-angle fault and along associated listric-normal and tear faults. The Bullfrog and Montgomery-Shoshone are volcanic-hosted, detachment fault-related, epithermal quartz calcite stockwork gold deposits. Bullfrog's minable reserve is 13 Mt (14.3 million st) of ore grading 3.8 g/t (0.11 oz per st) of gold. At Montgomery-Shoshone, 2.7 Mt (3 million st) minable tons at 2.5 g/t (0.072 oz per st) are present. Preproduction stripping has begun at the Bullfrog ore body. A 7.7-kt/d (8500-stpd) carbon-in-pulp (CIP) mill is under construction. In 1990, Bond Gold expects to be producing about 6.2 t/a (200,000 oz per year) of gold.

Jan 1, 1990

By Harold Wagner, Prakash Donde, Larry H. Heflin

INTRODUCTION The Washington Metropolitan Area Transit Authority is responsible for the planned 166 kilometer (103 mile) Metro rapid rail transit system in Washington, D.C. Some 117 kilometers (72.8 miles) of this system is in operation and contains 30.6 route-kilometers (19 miles) of rock tunneling. To complete the planned system an additional 1.6 kilometers (1.0 miles) of single track tunnels in rock are planned and additional contracts will consist of 4 kilometers (2.5 miles) of twin tunnel, resulting in 8 additional linear kilometers (5 miles) of single track tunnel contracts yet to be offered in soft ground. Three recent contracts in soft ground included the New Austrian Tunneling Method (NATM), along with conventional tunneling, as a bid option. Section E5a, Fort Totten Station and Tunnels, was the only one of the three to be constructed using the NATM option. This is the first application of the New Austrian Tunneling Method in soft ground in the United States. This paper will concentrate on the Fort Totten Station chamber NATM design, construction, and the instrumentation readings and evaluations which lead to an engineering back analysis which allowed a refinement of the bid design, resulting in a faster, more efficient construction. Design of the Fort Totten Station and Lines was done by Sheladia Associates. The design team included Greiner Engineering and Jenny Engineering with Mayreder-Consult of Austria, who was responsible for NATM design and NATM construction consultation.

Jan 1, 1991

By Ted M. Noskiewicz, James A. Ramsay

INTRODUCTION During the various stages of an underground project from inception through design and construction, probably the single factor of most interest and concern is the nature of the ground. The ground conditions more than any other item govern costs and feasibility; and although the methods available for assessing the ground have become increasingly reliable during the past thirty years, it is still very difficult in many cases to predict the behaviour of soil and rock during and after excavation. To help in this ground evaluation process, we have frequently supplemented the boring program with one or more large diameter inspection holes; and during the past eighteen years, seventeen of these holes have been constructed on tunnel projects which we have designed. There are references in the literature to similar inspection holes (1); however, as far as we are aware, no detailed description of them has been published. The behaviour of soil or rock during and after excavation and the effect of ground water on tunneling operations are of vital interest to tunnel designers and contractors. A comprehensive geotechnical investigation is therefore among the early and most important tasks carried out on a tunnel project and the backbone of every geotechnical investigation is the boring program. A series of borings, typically 50 to 100 mm (21 to 4 in) diameter, are drilled from which soil samples or rock core are recovered and the ground water observed.

Jan 1, 1979

By John T. Gormley, Pressley L. Campbell, Harry L. Crouse

INTRODUCTION This paper summarizes an investigation conducted to determine the relationship, if any, between reported occurrences of abnormal levels of carbon dioxide (C02) gas and the construction of a 2.44meter diameter main sewer tunnel under South Eighty-Fourth Street, West Allis, Milwaukee, Wisconsin. Specific objectives of the study were to determine: (1) the sources of C02, (2) the significance of each C02 source, (3) how C02 could move to areas where it was reported to occur in abnormal concentrations, and (4) what actions could be implemented to eliminate elevated C02 levels associated with the tunneling operation. To accomplish study objectives, expertise from several scientific and engineering disciplines was required including chemistry, geochemistry, microbiology, hydraulics, and hydrology. Pertinent documents were reviewed and field data pertaining to each aspect of the study were collected as required and examined to evaluate existing conditions. Background On or about February 1, 1971, gas with a composition different from normal atmospheric air was reported to enter basements and wells of residences in the vicinity of South Eighty-Fourth Street, West Allis, Wisconsin. Examination of the gas composition revealed that it possessed an elevated concentration of C02 and a depleted concentration of oxygen (02). Public health officials who investigated the appearance of the objectionable gas concluded within reasonable certainty that the elevated C02 concentrations were associated with the tunnel construction under Eighty-Fourth Street.

Jan 1, 1979

By C. E. Tackett

The Greenwich Collieries Co. is a subsidiary of the Pennsylvania Mines Corporation, and a wholly owned subsidiary of the PP&L Co. which serves the northeastern portion of Pa. PP&L annually consumes about 11 million tons of coal per year and is the largest consumer of both bituminous and anthracite coal on the eastern seaboard. Development work for Greenwich started in 1969. The coarse coal plant was constructed in 1970, with the raw fines bypassed. This worked very well for our requirements at that time; however, the Greenwich mining plans were extensively changed, longwall mining units were installed and low coal was encountered, resulting in dirtier coal, a finer coal and an ever increasing ash in the shipped product. All this demanded a more efficient ash removal process for the fine coal fraction. In 1971, the Pennsylvania Department of Environmental Resources and the Federal Environmental Protection Agency were pushing for the now stringent sulphur emission limits at our power plants. The Greenwich Mine coal reserves contained large quantities of 3.5 to 4% sulphur along with some low sulphur reserves. It was, therefore, essential that any new plant have the capability of efficient sulphur removal to meet the Pennsylvania Department of Environmental Resources and the Federal Environmental Protection Agency required limits. To meet these limits our design would have to be capable of producing a product of 2.0% sulphur or less.

Jan 1, 1977

By M. A. S. Martins, H. Medeiros, B. Y. B. Dos Santos

This paper contains the geometallurgical studies, performed for Mineração Rio do Norte bauxites, applied to a selected area belonging to Plateau Saracá III, referring to an exploitation period of one year and three months, approximately. It is worth mentioning that Mineração Rio do Norte has the biggest bauxite operation in Brazil producing 8,500,000 tons of concentrate per year. The studies included wide investigations directed to the final objective of establishing a precise mining and production planning, based on the different behavior of the distinct bauxite ores, when submitted to industrial processes. Geometallurgy can be defined as the study of the regionalization (space distribution) of the metallurgical variables of a mineral deposit. This study involves and integrates all process engineering, geology, geostatistics and mining planning, among others. The development and application of the geometallugical studies to Mineração Rio do Norte were motivated by several reasons and necessities, among them: 1. Improving the forecast of the amount of concentrate (metallurgical mass yield) produced; 2. Improving the equipment operational efficiency, providing a more homogeneous feed; 3. Conducting a more precise material balance that permits a better daily, monthly and yearly production control; 4. Improving the forecasts for the final products generated by each ore type; 5. Obtaining better knowledge of the behavior of several ore types, submitted to the different process unit operations; 6. Promoting adequate ore blending, producing an homogeneous feed during mining operation.

Jan 1, 1998

By D. E. Ranta, John G. Stone

One of the steps in an ore reserve estimate is customarily a rigorous calculation of the in situ grade of the orebody itself. The result of this calculation is usually reduced by some purely arbitrary factor to force the calculation to agree with the observed grade of ore actually mined. The failure of the calculation to realistically depict the grade of ore actually extracted is in many instances not the result of either an incorrect calculation procedure or of sloppy mining practice, but is a direct consequence of unsuspected irregularities in the contact between the orebody and the surrounding wall rock. A pattern of holes spaced at 1.5-3.Om foot intervals, drilled as part of a pre-development program, will often reveal that a presumed "regular' ore/waste contact is, in fact, so highly irregular that it will be next to impossible to separate ore from waste during mining. In such an instance, 'fingers' of waste projecting into the mining block will substitute for equal volumes of ore left outside the mining limits. As a result, the grade of material actually mined will be less than the grade envisioned for the orebody. The volume of waste substituting for ore in the mining block is about 0.4X, where X = the average error in projecting the ore/waste contact between alternate holes in the test pattern, and is expressed in units of volume per unit of area of the contact surface.

Jan 1, 1992

By N. S. P. Kumar

INTRODUCTION Development work at Carr Fork Mine of Anaconda Copper Company in Tooele, Utah commenced in August 1974. A decision to install an underground gyratory crusher for primary crushing run-of-mine ore and waste was made in 1976. Higher throughput rates, larger reduction ratios and greater flexibility in terms of meeting immediate and expanded production needs were the main reasons for selecting the gyratory crusher over the jaw crusher. UNDERGROUND MATERIALS HANDLING Underground crushing and conveying system was designed for a capacity of 16,300 mtpd of ore and waste on a 2-shift, 5-day week basis. Engineering and design of crusher complex was performed by Anaconda engineers at Carr Fork. Figure 1 represents the longitudinal section of mine showing the general materials handling system. Mine area is trackless except for materials and production haulage which is run on rail track on both 1200 and 970 levels. Ore from stopes and waste rock from development headings are handled by 3.8 cu m (5 cu yd) Jarco scoops and dumped into 2.4 m (8 ft) diameter ore and waste passes at the 1200 level. The material gravitates down to the 970 level where it is loaded into 12 cu m (420 cu ft) ASEA mine cars using side loading chutes. The muck train, consisting of 15 to 18 cars and two 27 mt General Electric trolley locomotives, is hauled over a distance of 2,300 m (7,550 ft) to

Jan 1, 1981

By Cherie M. Tilley

The Stillwater Mining Co. of Nye, MT, is operating a platinum-palladium mine that requires an extensive system of footwall laterals to facilitate mine production. The 10-year mine plan calls for more than 48,300 lineal m (160,000 lineal ft) of footwall lateral. A rapid tunneling and raising system is necessary to sustain mine production schedules and achieve production goals. This article describes the results of using a Robbins 4.1-m-diam (l3.5ft-diam) tunnel boring machine (TBM) to develop the footwall laterals and a Robbins RBM-6SP raise boring machine to develop raises at the mine. The Stillwater Mining Co is a partnership between Chevron Resources Co. and Manville Corp. Chevron Resources is the managing partner. When the mine began production in September 1986, the owners were primarily concerned with the actual production of platinum and palladium, the grade minable and the amount that could be mined. As mine development and mine production proved feasible, the owners considered higher production levels. The Stillwater Mine is developed by a system of adits (known as footwall laterals) that are driven parallel to the ore bearing zone. This led mine officials to consider using a mechanical boring machine in the primary mine development. Previously, tunnel boring machines have not been used successfully in mining applications. Therefore, the owners decided to purchase a used machine to reduce the potential risk.

Jan 1, 1991

By Theodore H. Budd, Douglas J. Heinz

This report concerns the successful construction of Contract No. 287 of the Northeast Side Relief Sewer System for the Metropolitan Sewerage District of the County of Milwaukee. The contract involved the construction of approximately 3316 meters (10,880 feet) of tunnel concrete lined to 1.83 meters (72 inch) internal diameter; 30.48 meters (100 feet) of 1.83 meters (72 inch) reinforced concrete pipe jacked-in-place; a river crossing of approximately 107 meters (350 feet); 197 meters (647 feet) of 1.83 meters (72 inch) reinforced concrete pipe in open-cut; several diversion and flow monitoring structures; and miscellaneous manholes and appurtenances. The contract proposal, which was extremely prescriptive, did allow alternative methods of primary support consisting of precast reinforced concrete segments, steel liner plates, steel ribs with wood lagging, and combinations thereof in designated areas. The specifications required the contractor to designate which support scheme it intended to use. The specifications also required compressed air as a means of ground water control for approximately 2316 meters (7600 feet) of tunnel. The alternative selected in the bid proposal was the support method consisting of steel ribs and wood lagging for 2263 meters (7425 feet) of tunnel and steel liner plates for the remaining section. Shaft supports, by specification, were designed to be steel liner plate with steel ribs varying in size from 4.27 meters to 6.10 meters (14 foot O.D. to 20 foot O.D. ).

Jan 1, 1983

By Kirsten Corin, Belinda McFadzean, Cyril O’Connor, Jenny Wiese

As orebodies being mined globally are of ever-decreasing grades, the focus on the concentration process has moved increasingly to the challenges faced in management of the gangue minerals. This is especially so in the flotation of platinum group minerals (PGMs), where head grades are typically in the region of 4–7 g/t. Hence, in PGM operations, a typical recovery of about 80 percent requires depressing about 85 percent of the gangue minerals. In a flotation cell the main control parameters are reagent type and dosage, froth depth and air addition rate. This paper reviews research projects carried out over an extended period aimed at optimizing the depression of gangue minerals: both those that are naturally floatable and those, such as chromite, that are entrained.  This has involved comparing the two major depressants used in the industry — guar and carboxymethylcellulose (CMC) — in the depression of the relevant gangue minerals, as well as investigating their consequent effect on the behavior of the froth phase. Because most PGM flotation plants treat the ore as if all the PGMs are associated with sulfides, copper sulfate is widely used as an activator, but this can also lead to inadvertent activation of gangue minerals. Finally, chromite levels must be kept below 3–5 percent to avoid any negative effects on smelter performance, and thus it is important to reduce chromite recovery which mainly occurs due to entrainment through, for example, froth depth control.   

By Todor P. Stefanov

The thermal draught within ventilation networks of underground mines depends to a great extent on the shafts depths and degree of change of ventilating air density. When the topology and space orientation of a ventilation network are constant, the dynamics of thermal draught is determined mainly by the variations of the heat- and mass-transfer processes in the openings. This conclusion follows from the well known basic theoretical notions about the thermal draught-aerostatics and thermodynamics [1]. Quantitative evaluation about the influence of heat- and mass-transfer processes on dynamics of thermal draught could be made through: direct mine measurements, special experiments and with the aid of mathematical models. The first two methods produce sufficiently accurative results, however it is very difficult to perform such studies, particularly when these processes are evaluated within a large scope and multifactor aspects. Mathematical models of thermal draught allow to overcome easily these difficulties. The degree of accuracy of results, obtained by this method, depends to a great extent on the degree of approximation achieved by the model to the real heat- and mass-transfer processes. In this paper, with the aid of Bulgarian developed method, algorithm and computer program V WT-4 - for complex evaluation of ventilation parameters in active mine ventilation networks [2], the authors investigate some changes in the thermal draught. Basically, the method is an iterating evaluation of an open set of equations, which represents relationships between the following basic parameters of a mine ventilation network: [ ]

Jan 1, 1980

By G. Deschenes, M. Fulton

The New Britannia Mine processes a qumtz-carbonate gold ore containing 4.63 git Au at a rate of 2,200 tpd. The ore contains approximately 3 .0% arsenopyrite, 2. 7% pyrite and the rest is gangue minerals. Eight to ten percent of the gold is refractory to leaching. Two years ago, New Britannia Mine experienced cyanidation problems. Gold extraction was not stable and occasionally decreased significantly. Preliminary test work, at laboratory scale and in the plant, indicated that lead nitrate addition was beneficial. A laboratory study was initiated to evaluate the level of control of the parmneters (grinding size, dissolved oxygen, pH, free cyanide and lead nitrate) to optimize the efficiency of the process. It was found that lead nitrate addition is required to alleviate the effect of metallic sulphides. This addition causes a significant increase in the gold leaching kinetics and a reduction in the cyanide consumption. The optimum system appeared to be: leaching for 24-30 hours at pH 11.8, 50 git lead nitrate, 250 ppm NaCN and 8 ppm 02. The gold extraction was 89.8% with grinding at 89%-74Fm (leached tail at 0.48 git). Cyanide and lime consumption are 0.40 kglt NaCN and 1.9 kglt respectively. Following the incorporation of lead nitrate, overall gold extraction was increased from 89.2% to 91.1 % but no change was noticed in cyanide consumption. The gross profit is $CDN 988,000 per year.

Jan 1, 1999

By G. Deschenes, P. J. H. Prud'homme

A laboratory study was carried out to evaluate the effect of oxygen and lead nitrate on the recovery of gold by cyanidation of a free milling gold ore, which contains 0.4 % copper as chalcopyrite. Efficient gold extraction is achieved only under specific conditions. The recovery of gold was increased from less than 90% to up to 98% by addition of lead nitrate. In the pre-leaching, the kinetics of oxidation of soluble sulphides are not an accurate indication for the length of the treatment. Even though the consumption of cyanide was reduced due to the inhibition of the dissolution of chalcopyrite by increasing additions of lead nitrate, cyanide consumption had to the held above 1.85 kg/t to maintain good recovery of gold. The continued addition of lead nitrate above 300 git had no impact on the recovery of gold. It was also found that the addition of lead nitrate at the start of cyanidation gave the same performance as when added during pre-leaching. The addition of oxygen increased the recovery of gold by 1.5%. The concentration ofNaCN had to be maintained in the range of700 mg/L because of the high concentration of copper in solution. The recovery of gold dropped significantly when the average NaCN concentration fell below 640 mg/L. Application of the findings of the laboratory study at the industrial scale led to improved plant performance.

Jan 1, 1997

By Byron R. Berger

Epithermal Ag-Au deposits form at low to moderate temperatures in near-surface environments. Historically, these deposits have been the source of high-grade ores. Increases in the market prices for gold and silver make lower-grade ores economically viable when available in large tonnages. Both the high-grade and bulk-tonnage ores are a consequence of the same processes, and general models of epithermal systems can be useful in the search for new precious-metal resources. A detailed study of deposits at Round Mountain, Nevada; Balei, Russia; Creede, Colorado; Guanajuato, Mexico; and Chanarcillo, Chile, indicates that the hydrodynamics of the fluid systems as determined by fracture geometries, the permeability, and composition of the host rocks are the key factors in the creation of both high-grade and bulk-tonnage epithermal deposits. End-member models that explain the range of observed deposit characteristics are: (1) the hot-springs deposition model; (2) the stacked-cell convection model (where mineral deposition occurs laterally along the interface of two partially separated fluid reservoirs); and, (3) the closed-cell convection model (where deposition occurs in vertical zones due to the relatively unimpeded vertical flow of hydrothermal fluids). Boiling and fluid mixing appear to be agents of precious-metal deposition with the self-sealing of the hydrothermal system being an important adjunct to boiling. Imaginative application of model concepts in exploration has led and can lead to the discovery of new high-grade and/or bulk-tonnage precious-metal deposits.

Jan 1, 1982

By Y. H. Alghamdi, A. D. S. Gillies

"INTRODUCTION Background Diesel Particulate Matter (DPM) has been a critical issue during the last decade in underground mines. This problem is threatening the health of mine workers, as well as the mine’s productivity. It has been confirmed that the diesel engine exhaust is the main source for the diesel particulate matter (DPM), and that it has the highest impact in raising carbon monoxide, carbon dioxide, oxide of nitrogen, and hydrocarbons in the underground environment. Miners who are exposed to this kind of emission are at high risk for developing lung cancer and other lung diseases. The Mine Safety and Health Administration (MSHA) has issued standard rules to regulate diesel emissions in underground mines. The final ruling confirmed in 2008 that the total carbon should not exceed 160 µg/m3. (MISHA, 2008) Monitoring DPM in underground mines has become important to ensure compliance with the MSHA regulations. There are many different types of DPM monitoring systems that are currently being used in underground mines. Two methods are typically used for DPM sampling: the personal sampling method or the direct exhaust sampling method, which takes a direct sample from the engine’s tailpipe. Many mines do not recommend using the direct exhaust sampling method since it involves complex measuring equipment and procedures, as well as it is not regulated in the United States or Canada. Thus, the personal sampling method is the preferred method in many mines when taking DPM samples. Even this method includes many variations, such as the Respirable Combustible Dust (RCD) method and the method devolved by the National Institute for Occupational Health and Safety NIOSH 5040. The NIOSH 5040 method is the preferred method overall because it gives more accurate measurements compared to the other methods. This method does not allow organic carbon sources other than those coming from diesel engine exhausts to interfere with the measurements, which otherwise would lead to overestimating the actual amount of the DPM in the mine environment. An impactor feature in the cassette, which holds the samples in the instrument, characterizes this method. Since the diesel particles are less than 0.9 micron, this impactor does not allow particles more than 0.9 microns to be collected in the cassette filter. The cassette is then analyzed in a lab and the concentrations of organic carbon (OC), elemental carbon (EC), and total carbon (TC) are determined. Lately, the National Institute for Occupational Safety and Health (NIOSH) has developed a new instrument that can use the same concept of NIOSH 5040, but with instant measurements of elemental carbon. It utilizes laser radiation with the cassette so it can give a real time measurement. The real time measurements give instant information about the underground environment that will allow the mining engineer to take immediate action and enact plans instead of waiting for results from the laboratories where NIOSH 5040 method may take several weeks to get the measurements."

Jan 1, 2016

By D. A. Gulley, L. J. MacDonnell, William A. Vogely

INTRODUCTION Environmental impacts are most commonly associated with air and water pollution. Not- withstanding this fact, to many people the primary disruption caused by mineral extraction is the change in the character of the locale. That is, the environmental impact of mining includes, but is not limited to, the generation of residuals. The dramatic transformation of wilderness or rural land often creates strong reactions to mineral development. Thus, the basic issues of land use management are a useful addition to environmental discussion. In this section, we take up a discussion of federal, state, and local land use management regimes, along with a brief introduction to the potential role of economic analysis. Elsewhere in this volume, land management issues are discussed from other orientations, and the interested reader may wish to look there as well. Locatable Minerals on Federal Lands Legislation and Administration: Access to hard-rock minerals on the public lands is governed by the General Mining Law of 1872. This law declares that "valuable mineral deposits in land belonging to the United States . . . are hereby declared to be free and open to exploration and purchase . . . ." In order to initiate rights, prospectors are required to physically stake claims of specified sizes and shapes following basic procedures set by federal law which may be supplemented by the individual states. Rights to these claims are protected at the exploration stage so long as actual possession is maintained and diligent exploration activity is underway. If a valuable mineral deposit is discovered on the claim, the claimholder earns the exclusive right of possession as against private parties so long as the claim is properly maintained. To maintain this exclusive right the claimholder must per- form $100 worth of labor or improvements, known as assessment work, on each claim every year and make related annual filings with the U. S. Department of the Interior. If the existence of a valuable mineral deposit can be demonstrated, and other legal requirements are met, the claimholder may receive a patent giving full title to the claimed land, including the surface in most cases. The purchase price is either $2.50 or $5 .OO per acre. The Mining Law has remained largely un- changed by legislative amendment during its now more than 100-vear existence. However, while the law itself has not changed, its interpretation and administration have changed considerably. Written at a time of rapid western expansion of the U.S. following the Civil War, the intent of the law was to promote mineral development by inviting unrestricted prospecting and by selling the lands containing mineral deposits at a minimal charge. As federal land policy shifted from one of rapid disposal into private ownership to one of retention and management for a number of uses, mineral patenting requirements have been administered more strictly. The most critical issue concerns what must be demonstrated to establish the existence of a valuable mineral deposit. The so-called "prudent man" standard was established in the case of Castle v. Womble (19 L.D. 455 (1894)):

Jan 1, 1985

By R. M. Rowe, A. K. Bhasin, D. Malhotra

Introduction Most primary molybdenite producers use a selective flotation process in which molybdenite is floated from the ore while other sulfide and nonsulfide gangue minerals are depressed in the same step. The process was developed at the Climax mill. Reagents used in the rougher flotation stage include vapor oil (collector), syntex (surfactant), pine oil (frother), and occasionally sodium cyanide and Nokes reagent (depressant). Rougher concentrate is upgraded by cleaner flotation to produce market-grade molybdenite concentrates. Reagents used in the cleaner flotation stage are Dowfroth 250, vapor oil, sodium cyanide, and Nokes reagent. The development of the flotation reagent scheme at Climax and the role of the reagents have been well described by Born et al. (1976) and Malhotra et al. (1976, 1977a, 1977b, 1977c, 1980). The hydrocarbon collector, vapor oil, accounts for approximately 50% of the total molybdenite flotation reagent cost. Because vapor oil sources are limited, and the cost of vapor oil has increased significantly over the last decade, several studies have been done to evaluate alternate oil-based collectors and/or blends of collectors that would produce similar metallurgical results at a reduced cost, or increase the overall molybdenite recovery at the same cost (Smit and Bhasin, 1983; Hahn, 1969; Smith and Bhappu, 1971). The approach taken in the present study was to evaluate alternate nonoil-based collectors and collector systems that are less expensive than vapor oil. A series of potential nonoil-based collector classes were identified and evaluated in the laboratory against the current collector system. The metallurgical performance of these collector classes was related to their chemical structures. The results of this study are discussed. Collector classes A survey of sulfide mineral collectors was performed with the objective of identifying and selecting potential collectors for molybdenite flotation. The available reagents were categorized into 10 general classes. Table 1 summarizes these collector classes, their chemical formulas, the manufacturers, and selected commercial collectors in each class that were evaluated in the present study. Experimental procedure A Climax ore sample, assaying 0.3% MoS2, 2.2% FeS2, 0.007% Cu, 0.003% Pb, and 0.03 oz per ton Ag, was used for this study. Standard laboratory rougher flotation test procedure involved grinding a 2-kg ore charge in a rod mill at 60% solids. Reagents were added to the rod mill including a collector, syntex (0.005 kg/t), pine oil (0.03 kg/t), lime (to adjust pH to 8), and sodium silicate (0.15 kg/t). The ground pulp was transferred to a 1000-g cell, and flotation was performed at 35% solids and 1800 rpm using a Denver D-1 flotation machine. Concentrate was collected for a flotation time of six minutes. The general approach used in the present study was to substitute the nonoil-based collectors at two to three different dosage levels (0.01 to 0.15 kg/t) for vapor oil in the bench-scale, open-circuit rougher flotation tests

Jan 1, 1987