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By Frank H. Madison

The percentage depletion deduction on Federal Income Tax Returns under the provisions of the 1939 Internal Revenue Code and of Sections 611 to Section 615, inclusive, of the revenue act of 1954, have created innumerable administrative difficulties, and a new interpretation of "gross income from the property" in that the Courts have expanded the ideas of what constitutes "gross income from the property" to a degree unheard of and unthought of prior to the enactment of these two codes. Originally, under the Revenue Acts of 1917, 1918, and 1920 depletion was allowable, based upon mine valuations as of March 1, 1913; if the mine was owned prior to March 1, 1913; upon cost, as of the date of acquisition if purchased subsequent to March 1, 1913; and upon a "discovery" basis if the mine was discovered after March 1, 1913, The "jab" for mine valuations as of March 1, 1913 and for discovery valuations and depletion was not as bad as it has been made out to be. Thousands of valuations and estimates of reserves were made and agreed upon between taxpayers and Internal Revenue Service engineers and as the years went on the answers have been found to be "reasonable." However, human nature being what it is each case was a battle of wits - of a sort - and many resulted in severe differences of opinion and litigation. As tax rates increased the situation was not bettered. I could talk for several hours on this phase of depletion but it is all past now andwould not give any real service or better understanding of the matter.

Jan 1, 1958

By Linda Gaines

Use of vehicles with electric drive, which could reduce our oil dependence, will depend on lithium-ion batteries. But is there enough lithium? Will we need to import it from a new cartel? Are there other materials with supply constraints? We project the maximum demand for lithium and other materials if electric-drive vehicles expanded their market share rapidly, estimating material demand per vehicle for four battery chemistries. Total demand for the United States is based on market shares from an Argonne scenario that reflects high demand for electric-drive vehicles, and total demand for the rest of the world is based on a similar International Energy Agency scenario. Total material demand is then compared to estimates of production and reserves, and the quantity that could be recovered by recycling, to evaluate the adequacy of supply. We also identify producing countries to examine potential dependencies on unstable regions or future cartels.

Jan 1, 2010

One of the major unknowns in the use of iron and copper is the quantity of various low level impurities, particularly hydrogen, that may be introduced during the deoxidizing process prior to casting. Introduction of hydrogen results in the embrittlement of components made from these metals. In addition, hydrogen bubbles in castings of copper result in the formation of blisters and swelling of the casting. During the last year, a significant effort has been made at the University of Missouri-Rolla to understand the permeation of hydrogen in galvanized automotive steel components and the hydrogen absorption during the electrodeposition of copper. The galvanized steel components consisted of automotive upholstery clips, seat belt holders and brake clips. Although the process of galvanization protects the steel from corrosion the process invariably causes the steel to pick up enough hydrogen to make it susceptible to brittle fracture.

Jan 1, 1992

By F. E. Joyce

This paper describes the development of a process for producing manganese and iron pellet concentrates from low-grade manganiferous iron ores of the Cuyuna range by a two-stage sulfatization-reduction roast. First, the manganese minerals (7.5 percent manganese) were converted to water-soluble manganese sulfate by roasting in sulfur oxide gases at 700° C, then, the hematitic iron minerals (31 percent iron) in the hot sulfatized are were reduced to magnetite with simulated reformed natural gas in the temperature range of 350° to 425° C. Batch tests were made in a 3-inch-ID by 6-inch-long, gas-heated, stainless steel retort, and continuous countercurrent experiments were performed in a 2-inch-ID by 10-foot-long, electrically heated shaft furnace. Manganese sulfate was removed from the roasted product by leaching with water, and over 90 percent of the manganese was recovered in an oxide product (65 percent manganese) obtained by precipitation, drying, and thelmal decomposition of the sulfate. An iron oxide concentrate (60 percent iron) with an iron recovery of 75 percent was obtained from the water-insoluble residue by magnetic separation.

Jan 1, 1973

By G. A. Perry

Cyprus Sierrita is a major producer of copper and molybdnum from its Sierrita and Twin Buttes open pit mines south of Tucson, Arizona. Both pits were started in the late 60's and are in excess of 1500 feet in depth. Sierrita has utilized a Cat 992C Remote Controlled Loader since January 1993. The loader was justified to increase SAFETY and PROFITS at Sierrita. It has exceeded all expectations. As of June 1993, it had produced additional profits totaling more than $1,500,000.

Jan 1, 1994

By B. Pollack, L. G. Stevens

INTRODUCTION In other chapters of this volume many of the important technical aspects in the design of concentration and dewatering circuits have been discussed. This chapter will deal with the steps necessary to bring that information together in order to produce the specifications necessary to purchase the equipment and complete the design. It deals with the plant as a whole rather than just concentration and dewatering, since the steps are essentially the same. Prerequisites to Engineering Design Feasibility studies will have been completed and design parameters established prior to the start of the engineering design stages for a metallurgical plant. The data base for the feasibility studies will have included laboratory and/or pilot plant test work, historical data, proposed plant site data such as climatology, topographic mapping, geological and geotechnical information. These studies will have established the general project design criteria as well as the process design criteria leading to the preparation of preliminary flowsheet ( s) and/or alternative process flowsheets. These will have incorporated metallurgical and material balances, allowing identification of major equipment items, and development of preliminary plant site plot plans. Preliminary technical and economic assessments and recommendations will have been prepared during this phase for Owner's review and evaluation of the technical, physical, and cost effectiveness of the overall project and its component process, process facilities, and supporting facilities elements. After Ownert s approval and decision to proceed, the engineering design can be initiated, using the above information as the basis. Engineering Design Phase The engineering design of metallurgical plants and their elements are generally developed in two sequential stages, namely: (1) Basic ("Front end") Engineering Design (2) Detail Engineering Design Figure 1 relates these engineering stages to the entire program from feasibility through construction.

Jan 1, 1986

By Arash Nikvar‑Hassani, Timothy Batchler, Lianyang Zhang

Recently, a new hybrid geopolymer/biopolymer (GP/BP) cementitious material was developed for improving the performance of pumpable roof supports in underground mines. This study demonstrates the application of the hybrid GP/BP cementitious material and validates its effectiveness in full-scale. In this regard, eight (8) full-size (0.61 m diameter and 1.52 m height) cribs were produced in collaboration with Minova International Ltd and tested at the National Institute for Occupational Safety and Health (NIOSH) Mine Roof Simulator (MRS) Laboratory. These full-size cribs were produced with different material configurations to evaluate the effect of water to solid (W/S) ratio, Portland cement (PC) content, and BP dosage. The results demonstrated and validated the effectiveness of the hybrid GP/BP cementitious material in increasing the peak and residual bearing capacities of pumpable cribs and eliminating the issue of deterioration when exposed to air compared with the conventional Portland cement/fly ash (PC/FA) cementitious material currently used in practice. On average, the peak uniaxial compressive strength (UCS) and the highest residual UCS after peak of the full-size cribs produced from the hybrid GP/BP cementitious material are 1.90 and 1.33 times of those of the PC/FA-based full-size cribs by one company and 2.32 and 1.66 times of those of the PC/FA based full-size cribs by the other company, respectively.

Feb 6, 2024

By M. Massinaei, M. R. Yarahmadi, S. Zeidabadi, M. Oliazadeh

"Although the majority of new flotation column installations continue to be for cleaning or recleaning, an increasing number of flotation projects are specifying columns for roughing or scavenging duties. NICICO has recently installed flotation columns operating in rougher, cleaner and re-cleaner circuits at three new copper concentrators; Miduk, Sungun and Sarcheshmeh. Miduk is probably one of the unique operations that use flotation columns in a rougher circuit. The Miduk rougher circuit comprises five CISA parallel columns. The rougher concentrate constitutes 88% of final concentrate and reports to concentrate thickener. The rougher tailings are scavenged with ten 50 m3 mechanical cells. The scavenger concentrate is then reground and cleaned in three 3.2 m columns in parallel. The cleaner tailings are scavenged in six 50 m37mechanical cells. The cleaner tailings along with the rougher scavenger constitute final tailings.In Sarcheshmeh, the cleaner concentrate is re-cleaned in a column. The column tailings are sent back to the cleaning circuit and the concentrate is reported to the concentrate thickener.This paper reviews performance details of the flotation columns utilized in rougher circuits at Miduk and compares their metallurgical performances in an industrial scale with columns at cleaning circuits for copper flotation. Practical limitations for applying the flotation columns in rougher circuits will be discussed in details.INTRODUCTIONThe use of flotation columns has grown rapidly since the first successful commercial application at Mines Gaspe, where columns were used in a cleaning circuit to float molybdenite. Early applications of columns were restricted to replacing or adding final cleaning stages in existing circuits. Columns can achieve superior cleaning as compared to conventional cells due to their ability to support a large froth depth and the use of water for washing froth. Wash water travels down through the froth and crosses the pulp/froth interface, replaces the entrained particles reporting to the froth from pulp and improve the concentrate grade."

Jan 1, 2011

By M. L. Bakker

Keywords: Pyrometallurgy, top submerged lance (TSL), nickel, PGM, calcium ferrite, lime ferrite The ISASMELT? top submerged lance (TSL) bath smelting process was developed in Mount Isa, Australia by Mount Isa Mines Limited (now a subsidiary of Xstrata plc) during the 1980s. By the end of 2011, the total installed capacity of the ISASMELT? technology will exceed 9 000 000 metric tons per year of feed materials in copper and lead smelters around the world. Commercial plants, operating in Belgium and Germany, are also batch converting copper materials in ISASMELT? furnaces. This TSL technology is equally effective for continuous converting processes, whereupon it is called ISACONVERT?. This paper presents the recently patented ISACONVERT? process for the continuous converting of nickel and Platinum Group Metal (PGM) containing mattes using the calcium ferrite slag system. The paper focuses on the potential application of the ISACONVERT? technology to existing nickel and PGM smelting complexes.

Jan 1, 2011

By Takefumi Takuma, Tsunenobu Nozaki, Masashi Nagano

The city of Naruto in the southwestern part of Japan is situated at the mouth of the mighty Yoshino River. An approximately 1,000m-long levee/bulkhead along one of the river’s tributaries in the city was raised with prefabricated steel panel walls on a concrete base bulkhead. Some sections of the walls’ foundations were combi-walls made of sheet pile cut-off walls and interspaced pipe piles; both of which were installed with press-in piling because the new wall alignment was very close to existing homes and businesses. A Clamp Crane, which clamped onto already installed piles and moved forward and backward on them, was utilized for a segment where the access to its piling location was too tight for a conventional crane or a track-mounted pile driver to maneuver.

Sep 8, 2021

By J. H. McKibbon

This paper outlines the basic steps in the sizing of electrostatic precipitators according to the Deutsch equation. Design parameters that do not appear in the equation are also described.

Jan 1, 1971

G. D. Kneip, Jr., and J. 0. Betterton, Jr. (Union Carbide & Carbon Corp., Oak Ridge, Tenn.)—The authors have contributed to the theory of zone melting by considering the effects of the solidification of the final zone on the distribution curves for the finite length bar. However, they do not consider in sufficient detail the restrictions imposed by the particular metals involved. Eq. 4 has the analytical solution which is equivalent to Pfann's' equation for normal solidification of the final zone. For distribution coefficients less than one, it is apparent from this form of the equation that, as the solidification approaches completion, the concentration attains extremely large values which are inconsistent with the density of physical materials. The necessary restrictions then are that the solute concentration cannot exceed the density of the alloy, or more frequently, the solubility limit in the solid phase. Secondly, the distribution coefficient cannot be constant from 0 to 100 pet solute unless the liquidus and solidus coincide over this whole region. It would thus be more correct to impose limits on the solubility such as would be indicated by a typical eutectic diagram. Furthermore, in this case the assumption that the distribution coefficient remains constant is more likely to be realized. Eq. 4 should then be replaced by the following C,,(x) =k/L-x [ ?11 3 C11 (x)Ddx - ?11 4 Cn(x)dx= k[Cn (L-l)/k] [ L-x/l]x-1 L-l = x = L-l [kCenterlie/Cn (L-l)]1/k-1 The typical shape for the distribution curve in the final zone is indicated schematically in Fig. 12 for the simple eutectic case. The solidification proceeds according to Pfann's' expression until the concentration reaches the maximum solubility in the solid phasc. At this point, the concentration changes abruptly to the eutectic composition and remains at this value for the duration of the solidification. The width of the flat region, and the back reflection of this effect into the concentration curve on subsequent passes, depends upon the ratio of the eutectic concentration to the original concentration, upon the distribution coefficient, and upon the number of passes. Similar effects on the concentration curves in the first zone length would be expected in many systems for which the distribution coefficient is greater than one, and where the maximum solubility is not too much larger than the original concentration. The discussers agree with the authors that the assumption that the liquid is uniform in concentration should be considered with caution. Kneip and Better-ton' have shown, however, that in floating zone refining of zirconium, using induction heating, the distribution of iron agrees with theory which assumes a uniform liquid composition. Hence, in this case, the experimentally realized distribution coefficient agrees with the present phase diagram within the experimental uncertainty. L. Burris, Jr., C. H. Stockman, and I. G. Dillon (authors' reply)—The authors are happy to have received the comments of G. D. Kneip, Jr., and J. 0. Betterton, Jr., which provide additional insight into the zone refining process. It is true that the equations developed are inapplicable if the solute solubility in the solid phase is exceeded. However, this restriction on the use of the equations was clearly stated in the paper. A closer approach to

Jan 1, 1957

By Simon Lekakh, Von Richards, C. Hank Rawlins, Kent Peaslee

"Batch aqueous leaching and carbonation tests were conducted using industrial steelmaking slags to determine the effect of carbonic anhydrase enzyme as a catalyst. Calcium leaching is a strong function of particle surface area, and the extent can be expressed as a function of time and particle size. Carbonic anhydrase did not affect the calcium-leaching rate, however, it did catalyze calcium carbonate formation to achieve a neutralization time near the theoretical rate. Additionally, carbonic anhydrase modified the precipitate morphology due to accelerated particle nucleation. Time controlled tests in which the pH dropped to ~6 decreased the amount of carbonate produced, and this effect was exaggerated by carbonic anhydrase, while pH controlled tests (>8.5) exhibited the highest rate of carbonation. Because the leaching rate was ~50% faster than the carbonation rate, a further increase in the amount of carbonation may be realized by using carbonic anhydrase however pH must be >10.3.IntroductionSteelmaking slag contains high fractions of alkaline earth oxide based phases (i.e., CaO and MgO) that exothermically form carbonates. Thus, it is being considered as a means of permanent carbon dioxide sequestration. At current U.S. steel production rates, approximately 13-17 million tons/year of slag are generated from basic oxygen furnace (BOF), electric arc furnace (EAF), and ladle metallurgy furnace (LMF) processes [1]. BOF and EAF slag is used as high quality mineral aggregate or cement clinker, while LMF slag, produced at ~15% the rate of BOF and EAF slags, has limited use as blast furnace flux or acid mine neutralization with the majority going to landfill [2]. Based on slag’s chemical nature and immediate availability as a co-product from steel production, the steel industry initiated a project to investigate hydrous carbonate formation in steelmaking slag as a method of sequestering carbon dioxide emitted in steelmaking offgas [3]. The goal of this project is to determine the reactor parameters suitable for the design of an industrial system for treating steelmaking offgas with raw or minimally processed slag."

Jan 1, 2009

Stawell Gold Mines Pty Ltd operates two underground decline mines within the township of Stawell. An active exploration program and commitment by owners to improving the ore reserves has led to an improved mine life and reduced operating costs since the purchase of the operation by Mining Project Investors and Pittston Mineral Ventures in late-1992. Over the past six years of operation a cumulative 1 150 000 ounces of gold have been added to the mineral inventory. An improved geological understanding of the orebody has evolved over time increasing the ore reserves and allowing the development of long-term plans for the mines future. This has enabled important capital decisions to be made providing the impetus for improved productivity and thereby reduced operating costs. Co-operative mine planning and mine geology utilising best practice geological systems have enabled the effective extraction of a structurally complex and mineralogically varied deposit within a rural Victorian township. Detailed metallurgical research has improved the function of an aged and poorly designed circuit to boost the mill capacity whilst also improving the metallurgical recovery. The combination of metallurgical expertise and an understanding of the geological setting enable an effective control of metallurgical recovery.

Jan 1, 2000

By H. Bartholomew, L. Bernstein

Experirmental bonds of Ag-In SLID to gold, copper, nickel, Kovar, Dumet, nickel-plated molybdenum, ALL-Pt vrzetallizing on ceramic, and nickel-plated Mo-Ti metallizing on Al2O3 ceramic have been made and exalnined. Metallographic taper sections were made of all bonds for Phase identification. In addition, failuve temperatltres were determined for bonds heated undev shear load. Sowe bonds were subjected to electvon-probe microanalysis. The ternary SLID system Ag-In-Sn was similarly evaluated. Plwtonticrogaphs and electron-probe data are discussed and related to over-all phase diagrams , diffusion effects, and wetting properties of the liquid. The rates of diffusion andor compound formation of gold, silver, and copper in indium and In-Sn liquids are found to descend in that order, as measured by the relatzve thickness of phase layers in the interfaces of bonds made at 200to 400°C. It is concluded that the SLID process enables one to produce bonds which are stable at high temperatures, after having been fabricated at low temperatures, and meet most requirements for application in interconnection, silicon bonding, or hermetic sealing of electronic devices. In a previous publication by one of the authors, the theory and principle of SLID bonding in simple binary systems was discussed. SLID bonding can be defined as a joining process using a preform with a laminated structure: a low-melting material clad on a high-melting core. During the bonding cycle a metallurgical transformation occurs such that the resultant structure can withstand, without reliquefying, temperatures higher than those at which the bond was made. The phases present in the joint interface region are high-melting compounds or, if equilibrium is reached, solid solution. Properties will depend on the phases present, ultimately approaching those of the solid substrate core of the SLID preform. The high-temperature strength of bonds made in these systems was correlated to phase identification in the interfacial bond regions from metallographic cross sections and electron-probe X-ray microanalysis data. When complex electronic components are assembled, multiphase systems result at the interfaces of bond regions. In these ternary and higher-order systems very little diffusion and phase-diagram data is available with which to predict the nature of the interfacia1 region that will result from conventional fabrication processes and subsequent field use of the component. Clarke and Rhines did early work in the field using the Al-Mg-Zn system as a vehicle. Guy and co-workers studied a portion of the Cu-Zn-Ni system and some mathematical4 and general5 aspects of multiphase diffusion. irkald' and coworkerse recently studied diffusion in the Cu-Zn-Sn system. No diffusion studies have been made on ternary systems directly applicable to this investigation but some phase diagrams are available: g-u-n, u-i-n, and CU-i-n. I) DESCRIPTION AND EXPERIMENTAL PROCEDURE The elements Au, Ag, Cu, Ni, In, and Sn are found in alloy systems commonly used in fabricating and assembling electronic components. Reference to the appropriate phase diagrams9 and the previous publication will indicate that the binary systems of any of the two low-melting elements, indium and tin, with any one of the other four higher-melting elements are capable of producing SLID bonds. It is not unreasonable to assume that any ternary or higher-order alloy system containing at least one of the elements gold, silver, copper, or nickel and one or both of the elements indium and tin could be used to produce bonds which are stable at high temperatures, after having been fabricated at low temperatures. Indeed, the liquidus contours and isothermal sections of those phase diagrams available7 do so indicate. The parameters affecting the bond and the equipment used in evaluating the bond were described in an earlier paper.' The current investigation used the

Jan 1, 1967

By N van Aarde, B A. Moore, S van der Merwe

Frequently materials handling operations are compromised by transfer chute problems ranging from high wear rates, blockages and ineffective loading of the downstream conveyor. This paper will present the design flow implemented to overcome problems in two case studies utilising discrete element modelling incorporating material characterisation.Material testing is relevant in achieving a valid discrete element method (DEM) simulation model and subsequently a valuable design tool for transfer chute assessment and design. It is well recognised that in the current growth and demand for mineral resources, leading to ever increasing conveyor capacities associated with export terminals and mineral processing facilities that transfer chutes are a key focus point. The successful performance of the transfer chute is also dependent on operating with a ærobust rangeÆ of material flowability and handling characteristics. This requirement is crucial in accommodating mineral ores from developing resources, which are often more friable and stickier than in past decades.This paper will include details of chute installations associated with iron ore and coking coal, from the DEM simulations through to the performance of the installed and commissioned plant.CITATION:Moore, B A, van der Merwe, S and van Aarde, N, 2013. Implementation of advanced transfer chute designs, in Proceedings Iron Ore 2013 , pp 445-450 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By Jamal Jafari, Abbas Sam

"The concentration of the fourth production line at Golgohar includes 66% iron with Blaine number of 500 cm2/g. For pelletizing process, the feed iron grade should be greater than 68% with minimum Blaine number of 1800 cm2/g. In this study, the low intensity magnetic Separators, particularly the feed particle size distribution of LIMS were considered. The feed particle size distribution of the separators was optimized to prepare an appropriate concentrate (as mill pelletizing feed). The samples were performed from the LIMS feed and after grinding, four d80 (64, 103, 132, and 15 3µm) were prepared. Then, laboratory magnetic separation experiments were conducted on these samples. The best results obtained using d80 of 132µm. In this particle fraction, the iron grade and Blaine number were 70.3% and 1571 cm2/g, respectively. Therefore, it seems necessary to include a grinding and classification systems before LIMS stage. According to the presence of a high amount of -132 µm particles in the initial feed, an inverse closed circuit was selected. The hydrocyclone was designed in three units as the classification system with 62 cm (diameter), 22 cm (overflow outlet) and 11 cm (apex diameter). By simulation of hydrocyclone with USIMPAC software, the size distribution of underflow was determined. Based on these results, a ball mill with dimensions of 6.7 m (length) and 5 m (diameter) was designed as grinding system.INTRODUCTIONGole-Gohar iron ore concentration plant No.4 (Polycom Plant) in Gole-Gohar Industrial and Mining Complex was installed to increase production capacity as well as treating the circulating load from AG and SAG mills of the main concentration plant of the complex. The load from the mine is crushed using jaw and cone crushers. The crushed ore is then combined with the circulated load and the combination is then grounded using HPGR and ball mills respectively.After comminution, there is a magnetic separation plant which includes a medium intensity magnetic separator (MIMS) followed by three low intensity magnetic separators (LIMS). Final concentrate dewatering is done using three belt filters. After the primary magnetic separation, the ore is upgraded to 62% Fe content and final concentrate which is the product of LIMS has 66% Fe content."

Jan 1, 2016

By E. M. Cox, A. S. Skapski, N. H. Nachtrieb, M. C. Bachelder

As a result of using copper-containing scrap in the steelmaking process, the copper content of steels has been steadily increasing for years. Consequently the possible role copper may play in the steelmaking process and in the finished product begins to attract the metallurgists' attention. Some time ago one of the present authors forwarded the idea—based on the results of the analysis of nonmetallic inclusions extracted electrolytically from steels—that sulphur in plain carbon steels is distributed mainly between copper and manganese, the amount of iron sulphide being very small; and that, consequently, the problem of copper and that of sulphur in steel cannot be treated separately.' At the time of the publication of the quoted paper little was known about the relative affinities of copper and manganese for sulphur at high temperatures except that at moderate temperatures (below 1000°C) the affinity of manganese for sulphur is much greater. To gather more experimental data on this subject, the present authors undertook the investigation of the equilibrium constants of the reactions: 2Mn(8 or 1) + S2(g) = 2MnS(s) 4Cu(s or 1) + S2(g) = 2Cu2S (S or I)* 2Fe(s) + S2(g) = 2FeS (s or 1) over a range of temperatures wide enough to establish the dependence of these equilibrium constants on temperature. From the equilibrium constants (K = l/Ps2) the free energy of formation (affinity) can be calculated from F° = -RTln 1/PSt (1) where the standard conditions chosen are: 1 atm of sulphur pressure and the activities of condensed components equal one. The decomposition pressure, Ps2, of sulphur over the respective sulphides is too small to be measured directly, but there is a way of eliminating this difficulty by measuring the equilibrium constant of the reaction between the sulphide and hydrogen. From the latter and from the equilibrium constant of the thermal dissociation of H2S we then calculate Ps2 for the respective sulphide. 2Mn + 2H2S = 2MnS + 2H, 2H2 + S2 = 2H2S_________ 2Mn + S2 = 2MnS The numerical values of the equilibrium constant of the thermal dissociation of H2S at different temperatures were taken from Kelley's paper, "The Thermodynamic Properties of Sulfur and its Inorganic Compounds."² In previous experimental work published by Jellinek and Zakowski3 and by Britzke and Kapustinsky4 the equilibrium constants of the reactions Metal sulphide + H2 = H2S + metal were determined by passing hydrogen, at different rates of flow, over the sulphide, analyzing the resulting H2S + H2 mixture and then extrapolating the H2S/H2 ratio (which is a function of the rate of flow) to the zero speed of flow, a method necessarily involving considerable uncertainty. In the present work the equilibrium ratio was actually measured instead of being extrapolated. The apparatus is shown in Fig 1. Experimental Procedure The sulphides were prepared by the following methods: FeS Powdered iron which had been reduced with hydrogen (ferrum reduc-tum) was mixed in stoichiometric ratio with sublimed sulphur and carefully ground. The mixture was put into an alundum crucible, covered with pure sulphur, and the reaction started by touching the mixture with a glowing iron rod. After the reaction was completed the product (still containing some metallic iron) was again ground with sulphur, put into a Rose crucible, covered with sulphur, and heated in a strong current of pure hydrogen. Analysis of the final product showed 62.46 pct Fe and 36.59 pct S. Theoretical for FeS: 63.53 pct Fe and 36.47 pct S. MnS Manganese sulphide (precipitated and carefully washed with distilled water containing H2S) was dried in a Rose crucible in an atmosphere of H2S and heated in a current of hydrogen for 2 hr at red heat. The product was then ground and ignited for several hours at 1000°C in a current of hydrogen sulphide. Analysis showed 64.53 pct Mn and 36.63 pct S. Theoretical: 63.15 pct Mn and 36.85 pct S. Some MnS samples were prepared from metallic manganese and sublimed sulphur by mixing and grinding them and then heating in a current of hydrogen sulphide in an alundum tube.

Jan 1, 1950

By J. F. Lupo

Leaching operations combine both metallurgical and hydraulic processes to extract metals from the ore heap. In the design of leaching operations, metallurgical tests and simple hydraulic tests are often conducted to define operational parameters such as leach cycle time, application rate, lift height, and effective leach area. While metallurgical constraints are often the primary focus on leach facility design, the hydraulic factors are also critical to the overall success of the operation. Some of the hydraulic factors that are known to impact solution flow through heaps include: • Ore physical and hydraulic properties, • Variation in ore properties under heap loading, • Ore mineralogy and its susceptibility to chemical and/or mechanical degradation, and • Heterogeneities within the heap. This paper presents a discussion on hydraulic factors that may impact the effectiveness of leaching operations. In addition, methods are presented which may be used to assess flow conditions within active and inactive heaps.

Jan 1, 2000

By Lyda W. Hersloff

An environmental monitoring program was established to assess the baseline of parameters which may be impacted by a research and development hydraulic borehole mining operation. During the baseline program, two aquifers were identified at the borehole site with the overlying sand aquifer being distinct from the ore sand aquifer on the basis of radiological parameters. The direction of flow of groundwater was determined to be north 19° west at a rate of 1.82 feet per year. Tests to determine the effect of backfilling the boreholes indicated that using overburden from the open pit mining operation would not significantly increase the radiological consti¬tuents of the ore sand aquifer. Mining measurements of the- radon flux from the borehole indicated that the actual release rate of 0.035 pCi/sec was several orders of magnitude less than the calculated estimate by the Environmental Protection Agency. Finally, a gamma survey conducted at the site indicated that the terrestrial component of background radiation was consistent with that in areas near uranium deposits. The methodology used as well as the supporting data are given in this report.

Jan 1, 1981