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By Bart Bartholomew, Randall J. Essex

INTRODUCTION In June of 2004, at the North American Tunneling Conference in Atlanta, an all-day workshop was conducted to review the application of Geotechnical Baseline Reports (GBRs) for underground construction. The moderator for the program was Bart Bartholomew, former President of Fruin Colnon Construction Company. The program involved commentary by a panel of engineers, owners, contractors, and legal representatives within the North American tunneling industry, as well as an extended audience discussion. The panelists were: Peter Douglass, President, Peter Douglass, Inc. Randall Essex, Executive Vice President, Hatch Mott MacDonald Joseph Gildner, Deputy Director for Technical Services, Sound Transit Randall Hafer, Esq., Kilpatrick Stockton Robert Fitzgerald Esq., Watt, Tieder, Hoffer, & Fitzgerald Kirk Samuelson, Senior Vice President, Kiewit Construction Group DEVELOPMENT OF PROGRAM FORMAT The focus of the program was to assess the state of the practice in the preparation of GBRs, and to discuss improvements that could and should be made with regard to their preparation, interpretation, and application. The approach used to achieve this focus in the broadest possible way was unique, in that it involved the panelists’ preparation of the workshop format and materials months before the presentation of the workshop. First, each of the six panelists was selected by the moderator based on their reputation and involvement in their respective fields of expertise, taking care to maintain balance and contrast within the collective pool of talent assembled. Following selection and agreement to serve on the panel, each panelist was asked to independently submit to the moderator a list of the six most pressing issues involving GBRs that they each frequently encountered in their individual professional experience. Each was asked to rank their issues in descending order of perceived importance. This process resulting in a pool of thirty-six potential or “candidate” issues for use in the workshop program. Contrary to the moderator’s expectations, there was surprisingly little duplication among the thirty-six issues submitted.

Jan 1, 2005

By John M. Dale

Sulfur has attractive features as a base material for coating systems. It melts at 246ºF and can be applied as a liquid above this temperature. Its melting point is just above the environmental stress level, 180°F, for most coatings and yet it is low enough to be easy to work with in comparison to many hot melt systems. The viscosity characteristics are such that sulfur lends itself to spray applications. The oxidation rate of sulfur in the atmosphere is very low and sulfur has outstanding chemical resistance characteristics, particularly in acid environments. One of the greatest potentials for these coatings appears to be in the protection of other materials. At one time, most acid tanks and pits were constructed with acid- proof brick and sulfur based mortars. This type of construction is much more expensive than sulfur based spray coatings. An attractive sulfur based coating composed of sulfur, talc, milled glass fibers and dicyclopentadiene,1 developed by the Institute for the Bureau of Mines has excellent acid assistance and can be used to protect concrete from acid attack.

Jan 10, 1973

By S. W. Parr

THE low-temperature carbonization of coal involves the carrying out of the coking process under conditions wherein neither the coal mass nor any of the passageways through which the volatile products pass are heated above 700° or 800° C. For convenience in this discussion, the single number 750° will be used to designate the maximum range. This temperature is not selected arbitrarily; it is the result of certain natural conditions that are inherent in the substances involved. Two of these conditions arc sufficiently pronounced to suggest a line of demarcation at this point as follows: (1) Below 750°, all the heavy hydrocarbons are expelled, which means that, at these lower temperatures, the illuminants, the gases of high calorific value, and the condensible oils are discharged; above 750°, there are given off the lean, non-illuminating gases consisting for the most part of hydrogen and marsh gas and having no condensible constituents present. (2) Below 750°, there is substantially no secondary decomposition; above 750°, the volatile products are readily decomposed, forming tars, naphthalene, free carbon, etc. It is not intended to maintain that no secondary decompositions occur below 750°. Many recent studies have demonstrated the practicability, especially in the presence of catalytic substances, of cracking certain of the hydrocarbon compounds; but at these lower temperatures the step is a moderate one, as, for example, from xylene to toluene or from toluene to an anthracene. These changes are moderate in amount. Not only do the reactions proceed slowly but they are subjected to the decomposing conditions for only a short time. This is evident when it is recalled that at these initial temperatures the decomposition, of the coal is very rapid and, if anywhere near a neutral pressure is maintained, the movement

Jan 2, 1920

By Kenneth K. Landes

Lake stone is a term currently applied to chemical and metallurgical grade limestone or dolomite that is brought down Lakes Michigan and Huron by large cargo vessels, mostly self-unloaders. This stone comes from 5 quarries, all in Michigan and all steel company owned. The volume in 1973 was over 34 million tons. The primary reasons for the large volume are the quantity and quality of the stone available, and the low delivery costs to lower lakes markets. The source rocks are Silurian and Devonian limestones and dolomites of exceptional purity. The regional dip of these formations is 50 to 60 feet per mile southerly toward the lowest point in the Michigan structural basin, which is near the center of the Southern Peninsula. Two of the 5 quarries, both high calcium limestone sources, are on Lake Huron in the northeastern part of the Southern Peninsula. The other three are in the Northern Peninsula; one at the northern end of Lake Michigan, and two on northern Lake Huron. The western quarry is in limestone, the other two in dolomite.

Jan 1, 1975

By J. A. Oraniscak

Longwall mining operations continue to increase extraction rates and the potential to generate higher amounts of airborne respirable dust (ARD). The National Institute for Occupational Safety and Health (NIOSH) has gathered data from underground longwall surveys and laboratory crushing tests to identify the important factors influencing the amount of ARD levels generated. Coal-seam characteristics, operational parameters (i.e.,cut sequence, mining height, etc.), extraction rates and dust-control parameters (i.e., airflow, water flow, etc.) were examined to determine their influence on ARD levels measured at longwalls during actual mining operations. ARD generated by a laboratory crusher while crushing bituminous coal samples collected from underground longwalls was also examined to determine if there were cor relations with coal-seam properties. Results from the longwall and laboratory data collected indicated that low-ash, high-volatile coals (low moist fuel ratio or MFR coals) generate higher amounts of ARD. It was found that face ventilation and water application to the coal product are essential elements to controlling longwall ARD levels.

Jan 1, 2000

By Thomas L. Kesler

IF all of the 14 million tons of kaolin produced in Georgia through 1949 had been mined from a single deposit 20 ft thick, it would represent a mined-out area of less than 1 sq mile. This measure of depletion is in such strong contrast with the large size of the area shown in Fig 1, from which most of the output has come, that there has developed an optimistic outlook concerning future supplies of kaolin equal in grade to that already mined. Dozens of mines are widely but not evenly distributed within this area, which consists mostly of parts of Twiggs, Wilkinson, and Washington Counties. There are scarcely half a dozen kaolin mines elsewhere in the state that have been appreciably productive, although the areas in which they occur may have good possibilities. Most of the efforts to find new deposits, therefore, will be confined to the area shown in Fig 1, for it is here that the industry has established its facilities, and prospecting is encouraged by relatively numerous outcrops and opened deposits of kaolin. The Ounulgee and Oconee Rivers drain the area. They and their main tributaries have flat, swampy flood plains covered with alluvium and for the most part thickly wooded. The divides between these streams are intricately dissected by the minor tributaries, which are spring-fed in steep-walled, cirquelike heads. Gullying and undercutting are active along the streams but are partly controlled by unusually dense vegetation. The mean altitude is about 350 ft above sea level and locally the relief ranges up to 200 ft. The area is near the inner margin of the Coastal Plain, and is underlain by Cretaceous and Tertiary formations. Its general location is shown in black in Fig 1. The division between the Coastal Plain and the Piedmont Plateau is actually a very sinuous line, the fall line, marking the extent to which erosion has stripped the Coastal Plain sediments from the southeastward- sloping surface of the crystalline rocks of the Piedmont.

Jan 1, 1952

By George Beavers

FLASH roasting of iron concentrate is in the experimental stage at Copperhill. The problem is peculiar in that the iron concentrate is predominantly pyrrhotite; the ratio of that mineral to pyrite being 5 to 1. Results so far indicate that the heat, generated in roasting a material containing 52.0 per cent iron and 38.0 per cent sulfur, requires special handling and for this reason the installation costs per ton of roaster feed may be too great to compete with hearth roasting. Interesting features of flash roasting at Copperhill are: 1. The sulfur dioxide content of the roaster exit gas can be varied between wide limits. Eight per cent sulfur dioxide gas, with 1.30 per cent sulfur in the calcine, has been common practice and gas up to 13.0 per cent sulfur dioxide has been possible. 2. Temperatures of the exit gases are very high. When roasting 60 tons or more of iron concentrate per 24 hr. in a modified Wedge roaster, the temperature in the roaster rises to 2000° F. and the calcine is lowered to 0.20 per cent sulfur. If the entire roaster system could be constructed to withstand such temperatures, ideal roasting conditions would obtain as regards sulfur elimination and tonnage roasted.

Jan 1, 1934

By J. Maheu

Heap leaching is an old technique for the recovery of copper, but a fairly new technique for the recovery of gold.The paper outlines the equipment required for the establishment of heap leaching test facilities and the procedure required to insure meaningful results. Sampling methods, the metal balance, carbon adsorption and stripping and safety procedures are discussed in broad outline.

Jan 1, 1982

By R. Moreau, V. Uspenski, K. Messadek, Y. Delannoy

Experiments and numerical simulations have been carried out to investigate the properties of quasi-2D turbulence. A new mercury model experiment has been built and tested in a uniform magnetic field up to 6T, (Ha~ 1000), ensuring a better two-dimensional behaviour than earlier versions (Alboussiere et al., 1997), limited to 0.2T. The quality of the experimental results has been improved in order to get wave number spectra (instead of frequency spectra) and higher signal over noise ratio (due to the higher magnetic fields). In parallel, some two-dimensional "direct" numerical simulations were performed using a commercially available code to solve a 2-D motion equation involving a source term to model the Hartmann layer. This term is derived from the condition of closure of the electric current, which allows to transfer the Hartmann effect into the 2D-core flow without simulating the layer itself (Sommeria & Moreau, 1982). No subgrid scale modelling is needed in such a two-dimensional situation where most of the energy is carried by the large scales.

Jan 1, 1999

By S Morrell, D Clark

æMine to MillÆ improvement involves optimising the chain of rock breakage processes from the in situ rock in the mine to the output of the concentrator. A growing number of paper studies and case histories from metalliferous mines (AusIMM, 1998; JKMRC, 1998; Clark and Scott, 2002) demonstrate that generating more breakage in the pit and the primary crusher can lead to benefits to milling that result in significant improvements to the economic performance of a mining project. However, unless the ævalueÆ of doing this (benefit over cost?) can be appropriately quantified, the approach cannot be effectively managed and the increases in blasting and monitoring costs cannot be justified.Mining operations have traditionally been organised into the separate departments of geology, mining and processing based on the specialist skills required and the different professional disciplines involved. Process optimisation across these boundaries has commonly been frustrated by management requirements to minimise the costs incurred by each department without adequate reference to the ævalueÆ being created. Companies have tended to discard positions that provide a technical over-view of the value chain, reducing the number of people with the opportunity to influence practices across these process boundaries.This paper discusses these issues and identifies measures required to quantify and manage the ævalueÆ provided by æMine to MillÆ improvement generated by the intense blasting of ore.

Jan 1, 2002

By T. R. Graham, R. G. Feustel, J. R. Long, R. S. Dean

The comprehensive study of the copper-manganese-zinc alloy system in the Bureau of Mines Laboratories has so far been principally concerned with alloys that lie within the alpha solid solution field of the system. Two regions, one centering about 65 per cent copper, 10 manganese and 25 zinc, and the other centering about 70 per cent copper, 20 manganese, 10 zinc, have been carefully examined. These have indicated that when the alloys are made with electrolytic manganese considerable amounts may be readily introduced and that the strength and hardness are increased without excessive losses in ductility. In the course of the work of establishing the limits of the alpha solid solution range,3 enough of certain of the alloys was prepared to permit determination of their- physical properties at a future date. The 60 per cent copper alloys concerned here are part of that series. They cross through the alpha field and into the adjacent alpha plus beta and alpha plus X fields, and are of interest because they offer some information on the effects of manganese on alloys lying in these fields. Composition and Fabrication of Alloys The chemical composition of the alloys involved is given in Table I. They were intended to contain 60 per cent copper with manganese increasing from 5 to 25 per cent in 5 per cent steps and zinc decreasing from 35 to 15 per cent. The analyses are so close to the intended composition that the alloys mill be referred to hereafter by their norninal compositions. They were finished into sheet by cold-rolling and intermediate annealing procedurzs calculated to produce sheet $i6 in. thick, cold-reduced by 20, qo, 60, and 80 per cent in thickness. Physical properties were determined on standard flat tensile specimens with 4-in. gauge length, although the test gauge section used was 2 in. The properties were obtained on cold-worked material and on material annealed at temperatures ranging from 700' to r500°F. after 60 per cent cold reduction. Since an annealing temperature of 1200°F.

Jan 1, 1946

By J. R. Long, T. R. Graham, R. G. Feustel, R. S. Dean

The comprehensive study of the copper-manganese-zinc alloy system in the Bureau of Mines Laboratories has so far been principally concerned with alloys that lie within the alpha solid solution field of the system. Two regions, one centering about 65 per cent copper, 10 manganese and 25 zinc, and the other centering about 70 per cent copper, 20 manganese, 10 zinc, have been carefully examined. These have indicated that when the alloys are made with electrolytic manganese considerable amounts may be readily introduced and that the strength and hardness are increased without excessive losses in ductility. In the course of the work of establishing the limits of the alpha solid solution range,3 enough of certain of the alloys was prepared to permit determination of their- physical properties at a future date. The 60 per cent copper alloys concerned here are part of that series. They cross through the alpha field and into the adjacent alpha plus beta and alpha plus X fields, and are of interest because they offer some information on the effects of manganese on alloys lying in these fields. Composition and Fabrication of Alloys The chemical composition of the alloys involved is given in Table I. They were intended to contain 60 per cent copper with manganese increasing from 5 to 25 per cent in 5 per cent steps and zinc decreasing from 35 to 15 per cent. The analyses are so close to the intended composition that the alloys mill be referred to hereafter by their norninal compositions. They were finished into sheet by cold-rolling and intermediate annealing procedurzs calculated to produce sheet $i6 in. thick, cold-reduced by 20, qo, 60, and 80 per cent in thickness. Physical properties were determined on standard flat tensile specimens with 4-in. gauge length, although the test gauge section used was 2 in. The properties were obtained on cold-worked material and on material annealed at temperatures ranging from 700' to r500°F. after 60 per cent cold reduction. Since an annealing temperature of 1200°F.

Jan 1, 1946

By John E. Sweet

There have already been so many subjects of a purely mechanical nature presented to the Institute of Mining Engineers, that it is unnecessary for me to apologize for adding another to the list. When a new invention, or what is supposed to be a new invention, claiming simply an improvement on the well-known devices, is to be described, it is usual to point out the defects of the old and give a history of the object or the art up to date. But, as no doubt a large proportion of the members of this Institute know as much, or more, about the history and use of the steam-engine indicator than I do myself, I will not further preface the subject, than to remind you that what called for the Richards's improvement on the Watt's indicator was high speed; that the leading improvement was the reduction of momentum in the reciprocating parts; and that this was accomplished by making the parts lighter, and reducing the travel of the heaviest, of them, and consequently the speed and momentum. The use of the indicator has been extended, for the speed of steamengines and other prime movers has been increased, until one engineer at least, John Cooper, of Philadelphia, says, "I want an indicator that will work up to a thousand revolutions." .Mr. Thompson, of the Buckeye Steam-Engine Company, of Salem, Ohio, has improved on the Richards's indicator by another reduction in the momentum, which is done by putting the parallel motion nearer the fulcrum of the pencil arm, reducing both weight and travel. The characteristic feature of the indicator I am going to illustrate is a still further reduction in the same direction, which is accomplished by dispensing with the parallel motion altogether. As is often the. case, in putting a simple idea into shape, some difficulties

Jan 1, 1879

By Michael D. McGill, Jeff Freadrich

There are many factors which affect the productivity of a front end loader. One of the most obvious is the characteristics of the material being mined. Characteristics such as size distribution, swell factor or heave, and material hardness / abrasiveness will all effect the efficiency of the mucking cycle. The focus of this paper is to look specifically at the size distribution of blasted material, and quantify the effects on loader productivity.

Jan 1, 1994

By H. P. Allen

The Achotla mill of the Cia. Minera de Penoles is of special interest in that it is one of the few still employing the chloridization process. The ore milled comes from the Suriana mine and consists of both oxide and sulphide, from neither of which can the gold or silver be extracted by a simple cyanide treatment. The mill was built in 1927, after several years of study in the development of the process1, and operated from December of that year until October 1930. Operations were resumed in February 1935, and have been continuous since, treating approximately 5000 tons per month, while the mine is getting up to full tonnage of 6500 tons per month. Transportation of Supplies The property is in Guerrero, Mexico, some 175 km. by airline southwest of Mexico City. The mill site is at Temisco (Fig. I), 8 km. from the Balsas River and 4.8 km. by aerial tram from the mine. At Temisco are the general office, shops, warehouses and laboratory serving both mine and mill. Here sufficient water is available either from a small creek running through the property or from the Balsas River, also convenient and ample storage space for the tailings. With the exception of the mine timber and a few native building materials which may be obtained locally, all supplies for both mine and mill are brought down the Balsas River from Balsas, the terminal of a Mexican National Railway line out of Mexico City. The method of transportation on the river is interesting. Rapids with very fast water are numerous, prohibiting the use of mechanically propelled boats. Those used resemble large, flat-bottomed rowboats some 22 ft. long by 8 ft. wide (Fig. 2). They are manned by two oarsmen sitting in the bow, each handling one large oar, and usually another man in the stern who, with his oar, assists in steering the boat, especially when the rapids are run. From 3 to 7 tons of freight are carried on each boat, depending upon the stage of the water, which governs also the time

Jan 1, 1939

By H. P. Allen

The Achotla mill of the Cia. Minera de Penoles is of special interest in that it is one of the few still employing the chloridization process. The ore milled comes from the Suriana mine and consists of both oxide and sulphide, from neither of which can the gold or silver be extracted by a simple cyanide treatment. The mill was built in 1927, after several years of study in the development of the process1, and operated from December of that year until October 1930. Operations were resumed in February 1935, and have been continuous since, treating approximately 5000 tons per month, while the mine is getting up to full tonnage of 6500 tons per month. Transportation of Supplies The property is in Guerrero, Mexico, some 175 km. by airline southwest of Mexico City. The mill site is at Temisco (Fig. I), 8 km. from the Balsas River and 4.8 km. by aerial tram from the mine. At Temisco are the general office, shops, warehouses and laboratory serving both mine and mill. Here sufficient water is available either from a small creek running through the property or from the Balsas River, also convenient and ample storage space for the tailings. With the exception of the mine timber and a few native building materials which may be obtained locally, all supplies for both mine and mill are brought down the Balsas River from Balsas, the terminal of a Mexican National Railway line out of Mexico City. The method of transportation on the river is interesting. Rapids with very fast water are numerous, prohibiting the use of mechanically propelled boats. Those used resemble large, flat-bottomed rowboats some 22 ft. long by 8 ft. wide (Fig. 2). They are manned by two oarsmen sitting in the bow, each handling one large oar, and usually another man in the stern who, with his oar, assists in steering the boat, especially when the rapids are run. From 3 to 7 tons of freight are carried on each boat, depending upon the stage of the water, which governs also the time

Jan 1, 1939

By K Nilsson, G Cheong

Troy Resources NL acquired a 70 per cent interest in the Mina Sertpo gold project in Goißs State, Brazil, in 2001 following a successful bid for the highly competitive international tender of a tenement package put out by Western Mining Corporation (WMC). The Sertpo deposit is located on cleared grazing land 380 km west of Brasilia and is 30 per cent owned by Amazonia Mineracao Ltda, a private Brazilian company. It is a small high grade open-pittable orebody with a measured and indicated resource of 108 000 tonnes at a grade of 49.7 g/t for 172 500 ounces using a 2.0 g/t cut-off. Troy undertook the project management and development of the project while Amazonia managed the exploration activities and provided other services as required. The project development involved drilling metallurgical test holes, scouring the world to locate an available and suitable second-hand plant, and arranging shipping to andreconstruction in Brazil.

Jan 1, 2004

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By E. Barnard

Tarkwa Gold Mine is located in south western Ghana, and is majority owned and operated by Gold Fields Ghana Limited. The Tarkwa CIL plant utilises a single low aspect 27' x 42' 14MW dual pinion drive SAG mill designed to mill 4.2mtpa (525tph). The plant was designed and constructed in a manner that would facilitate its expansion to 8.4mtpa (1,050tph). Since the plant was commissioned in September 2004, the SAG mill has consistently exceeded its nameplate throughput by up to 25%. Through the installation of a 26? x 36? 14MW dual pinion drive Ball mill along with additional upstream and downstream equipment, the plant will now be expanded to 12mtpa (1,500tph). This paper endeavours to outline the process and engineering design challenges along with the future construction challenges of the expansion in the midst of a fully operational plant.

Jan 1, 2008

By O. Mellgren, A. M. Gaudin, P. L. De Bruyn

The adsorption density of ethyl xanthate on pyrite was determined as a function of xanthate concentration. Surface preparation of the mineral appears to have asafunctionsome effect on the subsequent adsorption process, A monolayer of xanthate on the surface is exceeded only in presence of oxygen. The effect of OH- , HS- (and x and CN- S=)and on the amount of xanthate adsorbed was investigated. Competition between OH- and X- (xanthate) ions for specific adsorption sites is indicated over a wide pH range. IN the flotation of sulfide ores, xanthates are most commonly used to prepare the surface of the mineral to be floated so that attachment to air takes place. The quantity of agent required to make the mineral hydrophobic is usually very small, of the order of 0.1 to 0.25 lb per ton of mineral. Details of the mechanism of pyrite collection are for the most part unsettled. Adsorption of collector has long been believed to involve an ion exchange mechanism as demonstrated for galena' and for chalcocite.2 In the work on chal-cocite it was also demonstrated that a film of xanthate radicals unleachable in solvents that dissolve alkali xanthates, copper xanthate, or dixanthogen was formed at the surface of the mineral. The unleachable product increased with increasing addition of xanthate up to a maximum corresponding to an oriented monolayer of xanthate radicals. Pyrite is extremely floatable with xanthate if its surface is fresh.9 ut the floatability decreases rapidly as oxide coatings increase in abundance. Pyrite shows zero contact angle when in contact with ethyl xanthate solution at pH higher than about 10.5;4 at neutrality, a contact angle of 60" is obtained at a reagent concentration of 25 mg per liter. Alkali sulfides and cyanides are pyrite depressants. In this study of pyrite collection the writers have sought to relate measured xanthate adsorption to the method used in preparing pyrite, to the presence or absence of oxygen, to concentration of hydroxyl, hydrosulfide, sulfide, and cyanide ions. The principal experimental tool has been radioanalysis," " using xanthatcx marked with sulfur 35. Experimental Materials Pyrite: Unlike most sulfides, pyrite is a poly-sulfide. The structure given by Bragg7 resembles that of sodium chloride, the iron atoms corresponding to the position of sodium and pairs of sulfur atoms corresponding to the position of chlorine. The edge of the unit cell in pyrite is 5.40 A and in halite 5.63 A. The S-S distance in pyrite is 2.10 A; the Fe-S distance, 3.50 A: and the Fe-Fe distance, 3.82 A. Natural pyrite from Park City, Utah, was used in this investigation. Pyrite 1 was obtained by hand picking pure crystals. Pyrite 2 and Pyrite 3 were obtained from finer textured crystalline material containing inclusions of silicates. The same cleaning technique was utilized for the preparation of Pyrite 2 and Pyrite 3, whereas a different cleaning technique was used for Pyrite 1. Pyrite 1 was prepared as follows: The crystals were ground in a porcelain ball mill and the 200/400 mesh fraction was separated by wet screening with distilled water, followed by washing for 1 hr with deoxygenated distilled water acidified with sulfuric acid to pH 1.5. The acid was removed by rinsing with deoxygenated distilled water on a filter until a pH of 6.0 was reached in the effluent. This filtration was carried out under nitrogen. The sample was then dried in a desiccator under nitrogen. The period of time for which this pyrite sample was in contact with water containing oxygen was about 4 hr. The specific surface as determined by the BET gas adsorption method was 582 cm2 per g. Final material assayed 53.12 pct sulfur and 46.5 pct iron (theoretical, for FeS,: S, 53.45 pct; Fe, 46.55 pct). After crushing, Pyrite 2 and Pyrite 3 were washed with 1 M HCl. rinsed, and fed to a laboratory shakinq table to remove the small amount of silicates. The concentrate obtained was ground in a laboratory steel ball mill. The 200/400 mesh fraction was separated by classification in a Richards hindered settling tube. This fraction was then given a final wash with 0.1 M HCl and deoxygenated water was filtered through the sample. The final effluent showed a conductivity equivalent to that of a solution having a salt concentration of 0.3 ppm. Aqueous hydrogen sulfide solution was then added to the sampln (about 100 ml saturated H,S solution to about 1000 g pyrite under a few hundred milliliters of water) which was stored wet under nitrogen. The sample stored in this manner showed no indication of formation of iron oxides, whereas iron oxides appeared

Jan 1, 1957