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By Robert S. Sanford

In April 1948, the Bureau of Mines was requested to evaluate the graphite deposit' known as the Benjamin Franklin and Just mines in Chester County, Pa. After preliminary investigation, it was recommended that the Bureau core-drill, trench, and sample the deposit and make beneficiation tests on the ore. During the first and second World Wars, flake graphite was in short supply, and a number of deposits in the United States were exploited. In 1942, the Defense Plant Corporation, through its lessee, the Benjamin Franklin Graphite Co., purchased 70.263 acres of graphite-bearing land and built a complete mining and milling plant with a capacity of 300 tons of crude ore per 24-hour day at a cost of $426,250 (Plancor No. 1254).

Jan 1, 1949

By Alton G. Woody

Like so many installations of its type approaching middle age, the ore processing and shipping facilities in Puerto Ordaz, Venezuela, have taken a lot of twists and turns in their lifetime. The object of this paper is to lead you through these twists and turns and some of the reasoning for them. The need for additional sources of iron are in the United States became evident during the time of the second world war. An exploration team from U. S. Steel located further high quality direct shipping iron are in commercial quantities near Ciudad Bolivar, Venezuela in 1947. Basic decisions were made to develop these reserves for export via the Orinoco River through a new shipping port to be built at the confluence of the Orinoco and Caroni Rivers. The basic concept was for the ore to be loaded directly to ocean going vessels. This was something of a gamble since the basic economics depended upon successfully opening up the main channel of the Orinoco River to deepwater vessels, and the Bethlehem Steel mine in the same area was already working with shallow draft river boats and a transfer station on the coast. Always, the basic philosophy has tended to be to plan with expansion in mind. The original conveyor systems were 1.52 m (60 inches) wide with 200 idlers, whereas 1.07 m (42 inches) or 1.22 m (48 inches) wide conveyors could have carried any production foreseen at the time of installation. Stockpiles have been located and designed with room for expansion. Basic designs have had dotted line future expansion blocked in.

Jan 1, 1978

By M C. Ackland

The Grande Côte Mineral Sands Deposit (GCMSP) is located on a coastal mobile dune system in Senegal on the west coast of Africa. Senegal is a stable, democratic republic, having gained its independence from France in 1960. It has a population of about 12.5 million people. The dunes begin about 25 km north-east of Dakar, the capital city of Senegal, and extend northwards for more than 140 km. The mineralised dune system averages about 2 km wide with some areas extending to up to 4.5 km wide.Mineral Deposits limited (MDL), a Melbourne-based company with over 60 years operational experience in mineral sand mining in Australia, holds, through its Senegalese subsidiary MDL Senegal SARL (MDLS), a 25-year Mining Concession covering the deposits. The main heavy mineral (HM) areas identified to date are, from south to north, Mboro, Fass Boye, Diogo and Lompoul. Within the Mining Concession there is potential to identify additional resources beyond the limits of present drilling. The dunes are mobile or semi fixed, pale yellow in colour and overlie older white marine sands. The dunes range between 5 m and 30 m in height and the mineralised zones, which are essentially flat-lying, average around 15 m in thickness. The mineralisation consists, principally of ilmenite with accessory zircon, rutile and leucoxene. Zircon and ilmenite are the main commodities of interest.Independent consultants AMC estimated mineral resources for the four main deposits in 2006 of 1330 Mt of sand averaging 2.0 per cent HM at a cut-off grade of 1.5 per cent HM. At a cut-off grade of one per cent the resources were 3200 Mt averaging 1.5 per cent HM. All of the estimate was classified as Inferred Resources under the 2004 Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (JORC, 2004).Subsequent completion of infill drilling, sampling and assaying of the initial mining area, centred at Diogo, was completed in 2008, enabling a final dredge path to be designed and ore reserves estimated, based on a JORC compliant Measured and Indicated Mineral Resource. Infill drilling is continuing in the Lompoul and the results of this campaign will be included in the JORC compliant data base by AMC in April 2010 to enable the 15 per cent year dredge path and mine plan to be finalised.Development studies have shown that the project can be optimised as a dredging operation, with mineralised sand being treated in a conventional floating spiral preconcentrator, wet concentrator and dry mineral separation plant. Engineering work now substantially completed by Ausenco Limited (Ausenco) has developed the combination of some existing plant from MDL’s past operations, and new plant for the complete process from dredging to final product separation and grading. Testwork on a series of bulk samples has determined that the project can yield a high quality zircon product and an ilmenite product along with minor rutile and leucoxene.Work to date indicates a likely output of around 80 000 t of zircon in various quality streams (Primary, Secondary, Standard and Foundry Grades), around 550 000 t of ilmenite plus around 15000 t of rutile and around 12 000 t of leucoxene/secondary ilmenite. Planned zircon output would represent around 7per cent of the total world production, which would make the GCZ&IP a significant producer on a world scale.The studies have indicated capital costs to be in the range of US$400 M total operating cost (after completion of ramp-up) of around US$50 M - $55 M/a. Revenue is expected to be in the range $127 M - $157 M/a.

Oct 5, 2011

By Wilson W. Zoogah, Joel G. Burken, Abdullah Al Moinee, Mariam K. Al-Lami, Kwame Awuah-Offei

Soil CO2 flux serves a critical indicator of ecosystem function and carbon cycling in disturbed landscapes. This study measured and quantified temporal patterns and environmental drivers of soil respiration across fifteen revegetated treatment plots over a 12-week period (July-October 2025). Utilizing automated CO2 flux measurements, we investigated differences between treatments and quantified their relationships with soil temperature, moisture, and atmospheric pressure. One-way ANOVA revealed significant treatment effects on CO2 flux (F4, 442 =14.1, p < 0.001, η2=0.113), with amended plots exhibiting 40–50% higher mean fluxes than control plots. Weekly temporal analysis demonstrated distinct temporal patterns, with peak fluxes registering in mid-August before a gradual decline. Multiple regression analyses showed that soil temperature (β=0.043, p < 0.001) and moisture (β=2.31, p < 0.001) are the primary flux drivers, together accounting for 18.2% of total variability (adjusted R2=0.182). Atmospheric pressure had a moderate effect (β=1.13, p < 0.001), detectable only in multivariable analysis. Variance partitioning indicated that environmental factors explained more variability (18.7%) than treatment effect (10.5%), with approximately 71% of flux variability attributed to finer spatial or temporal variability. These results demonstrate that revegetation amendments enhance soil biological activity and impact geochemical cycling, and that temperature and moisture are dominant controls on respiration dynamics in reclaimed mine soils. The findings provide quantifiable benchmarks for ecosystem recovery through altering soil biology and geochemistry and for quantifying carbon sequestration potential in post-mining landscapes.

Feb 22, 2026

By E. O. Binyone

"https://www.youtube.com/watch?v=PBPyGXjlBHIhttps://www.youtube.com/watch?v=3AbtiKOYg7Y&feature=push-fr&attr_tag=uKtVW3yq2ZngptnE%3A6https://www.youtube.com/watch?v=PR2THayf-N8https://www.youtube.com/watch?v=mF-BXiWocqIhttps://www.youtube.com/watch?v=TuesIw-OZA4https://www.youtube.com/watch?v=v2HP7sTIZgEhttps://www.youtube.com/watch?v=v2HP7sTIZgEINTRODUCTIONIn 1937 the Eureka Corporation, Ltd., of Canada obtained a lease on the mining property of the Richmond-Euroka Mining Co. in the Eureka mining district, Eureka County, Nevada. In 1910 the Richmond-Eureka Co. suspended all operations except small-scale leasing, but company and other geologists contended that the El Dorado dolomite would also carry ore in its continuation from fault block north and northeast of the productive old working. The company began to explore the fault block in 1923 by diamond drilling from the,900 level of the Locan shaft, but for various reasons the hole was drilled to a depth of only 2 feet and was then abandoned. Eureka Corporation, Ltd., planned an extensive diamond-drilling program, which included four deep diamond-drill holes, one designated ""A"" fig. if it is to be drilled from the surfaced, and three designated ""E"", ""C"" and ""D"" figs land to be drilled from the 900 level of the Locan shaft. Of these only B and C were completed by the company. Hole C was a contribution,of the hole started by the, started by the Richmond Eureka Co. The site for hole D (figs. 1 and 2) was prepared, but the hole was not drilled. Both B and C holes penetrated sulfide ore carrying good percentages of zinc, lead, gold, and silver. The ore bodies were 40 feet thick and were found approximately 1,400 feet below the 900 level of Locan shaft.On the basis of the very favorable results obtained from the drilling program, the Eureka Corporation planned extensive development of the property and in 1942 commenced sinking the Fad shaft (figs. 1 and 2). The shortage of equipment and materials brought on by the war and difficulties with water in the shaft greatly hindered progress. However, early in 1943 the Eureka Cor¬poration, Ltd., of Canada submitted to the War Production Board a proposal for developing the mine and milling 500 tons of zinc-lead ore a day. The proposal involved an expenditure of $2,890,750."

Dec 1, 1946

By Y Xie, L O’Connor, S C. Dominy

"Sheeted vein gold deposits are characterised by multiple subparallel veins and free milling gold. In situ heterogeneity of mineralisation results in grade and process parameter variability, which will increase project risk. Measured variability is often enhanced by poorly designed sampling and test work protocols. Protocols that are optimised within the framework of the theory of sampling to suit the ore type, together with quality assurance/quality control systems, will reduce variability and provide fit-for-purpose results. The geometallurgical approach requires spatially distributed samples within a deposit to support variability modelling. Diverse attributes from core logging, mineralogical/textural determination and small-scale tests are used to measure variability. This contribution emphasises early stage geometallurgical intervention during project development. It presents a discussion of representative sampling in the context of the theory of sampling. A case study presents a protocol that exemplifies how data can be gained from a well-designed and planned program to support a resource estimate and prefeasibility study.CITATION:Dominy, S C, O’Connor, L and Xie, Y, 2016. Sampling and Test Work Protocol Development for Geometallurgical Characterisation of a Sheeted Vein Gold Deposit, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 97–112 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By T. W. Smelser

This Bureau of Mines report presents a review of underground testing, test results, and design guidelines of steel-fiber-reinforced concrete (SFC) mine support (cribbing) developed at the Spokane Research Center. The objective of the work was to verify, in field applications, the performance, economics, and acceptance by the mining industry of the SFC supports and establish design guidelines for their use underground. Results indicate superior performance, favorable economics, and enthusiastic acceptance by the mining industry.

Jan 1, 1983

By S C. Dominy, R C. A Minnitt

Determination of the sampling constant and liberation diameter are of importance when applying the Fundamental Sampling Error (FSE) equation to optimise sampling protocols. The method for the direct determination of the sampling constant is based on the æGy 50-pieceÆ experiment and variants thereof, which is generally known as the heterogeneity test. In the coarse gold environment, the heterogeneity test is prone to poor repeatability and bias due to the very heterogeneity that it aims to quantify. A key issue is the large sample mass that can be required in the coarse gold environment, which may reach 1 t or more. The heterogeneity test often provides an evaluation of the background gold grade heterogeneity, but understates that of the high-grade coarse-gold component. This is because the total mass of fragments is usually too small to represent the mineralisation. Despite this relatively well-known problem, heterogeneity tests are often undertaken on deposits where the presence of coarse gold is ignored, not fully realised or vastly understated. Sampling types and protocol designs are frequently based on poor quality data that do not match the characteristics of the orebody. For this style of mineralisation, an empirical approach to define ore characteristics is recommended.CITATION:Dominy, S C and Minnitt, R C A, 2012. Application of heterogeneity testing to coarse gold ores, in Proceedings Sampling 2012 , pp 143-160 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 21, 2012

By Allen L. Frank

The uranium mine environment is characterized by airborne radon-222 and its daughter products to which miners are exposed through breathing and by gamma and X rays from the uranium-235 and uranium 238 chains to which the miners are exposed externally. The nature of the radon-daughter concentrations and the methods used to define and measure the exposures are explained through an extensive review of the literature. An improved passive track-etch dosimeter for personal use is proposed which has a measurement range of about 30-400 WLH with a 95% confidence interval to ~25% for individual detector measurements in relatively stable mine air. Some measurements of these dosimeters adapted to environmental use are also presented. A program of development for enhancement of dosimeter sensitivity and greater accuracy of measurement is outlined. Extensive gamma-ray survey measurements were made in two Colorado and three New Mexico mines. Follow-up personal miner dosimetry with LiF TLDs was done at the mine with the highest exposure rates, The average miner exposure rate measured was 4.27 mR per day, or 9.3 of the maximum permissible level averaged over 13 weeks. Comparisons are made with previous mine surveys and their value in estimating miner exposure is discussed. Some depth-dose measurements made using Lucite as the absorber are evaluated.

Jan 1, 1978

By I. Nishkov, I. Grigorova, M. Ranchev, S. Dzhamyarov

"Chelopech concentrator (part of Dundee Precious Metals) is located in central-western Bulgaria. The current process plant utilizes conventional primary crushing and SAG mill grinding, copper and pyrite rougher and cleaner flotation stages to produce a marketable gold-rich copper and pyrite concentrates. As in the most processing plants, the pyrite depression is carried out alkaline media using lime. The subsequent activation of pyrite mineral is achieved by adding sulfuric acid, which has an adverse effect on the flotation equipment. Therefore, a series of laboratory flotation experiment for activation of lime depressed pyrite by ammonium salts, such as ammonium chloride (NH4Cl), ammonium sulfate (NH4)2SO4 and ammonium bicarbonate (NH4HCO3), have been conducted. The obtained results showed that ammonium bicarbonate proved to be the most appropriate reagent choice for pyrite mineral activation. These results are confirmed by X-ray photoelectron spectroscopy (XPS), performed to determine the surface chemistry of pyrite in its natural state, after lime treatment and subsequent treatment by NH4HCO3. The registered XPS spectra indicate that the hydrophilic phases on the pyrite surface could be efficiently removed by using the ammonium salt NH4HCO3. The industrial use of NH4HCO3 as an activator of lime-depressed pyrite is a possible alternative for sulfuric acid replacement as a most widely used pyrite activator. Flotation experiments with addition of 200, 400, 600 and 800 g/t from each of the studied ammonium salts have been conducted. Furthermore, the impact of the flotation pulp pH within the scope of 9.00 – 10.50 was also investigated. The obtained results from the laboratory flotation experiments, showed that the pyrite activation with ammonium bicarbonate is an excellent possibility for replacing the sulfuric acid as main pyrite activator reagent, in order to eliminate the gypsum accumulation during the flotation of pyrite in Chelopech concentrator.INTRODUCTION The Chelopech deposit, located about 60 km east of Sofia is one of the largest Cu-Au deposits in Europe. The ore have complicated mineralogical and chemical composition. The main ore minerals are represented by pyrite, enargite (luzonite), tennantite, chalcopyrite and bornite. The average copper content of the mined ore is about 1.1 % and the gold content is about 4.0 g/t."

Jan 1, 2018

By H. S. Castro

The by-product molybdenite technology in Chile is reviewed. Flowsheets, reagents and some operation features for the actual six Chilean moly plants(Chuquicamata, El Teniente, El Salvador, Andina, Las Tortolas and Pelambres), are presented. The chemical and physicochemical factors involved in the depression of copper and iron sulfides during the selective flotation of molybdenite, with sodium sulfide, sodium sulfhydrate, anamol-D and LR-744, are discussed. Advances in technology to solve problems of excess of lime (precipitation with sulfuric acid); excess of residual flotation reagents (activated carbon and kerosene); and optimization of depressant consumption (Eh control, nitrogen and low-oxygen air), are discussed.

Jan 1, 1998

By B. Das

The separation of iron values from a typical Indian iron ore slimes containing 57.2% Fe, 5.81% SiO2 and 7.05% Al2O3 by wet magnetic separation and selective magnetic coating is investigated. The iron phases, which are mostly paramagnetic in nature, have been coated with magnetite or colloidal magnetite in presence of surfactant by manipulating the surface properties of the minerals. The studies have indicated that an iron concentrate containing 62.32% Fe with 66.5% iron recovery can be obtained at medium magnetic intensity of 7.8 kG which is much less compared to normal magnetic intensity of 13 kG to14 kG required for hematite-goethite separation. The performance of this technique with respect to both grade and recovery is superior to conventional magnetic separation or flotation of iron ore slimes.

Jan 1, 2014

By F. W. Wessel

Manganiferous carbonate slates of the Cuyuna range, which are part of the unoxidized thin-budded iron formation, represent a large potential source of manganese. This report details recent bench-scale research by the Federal Bureau of Mines in applying mineral-dressing techniques to these materials in order to investigate their potential uses. Work outlined in the report was directed toward the production of concentrates usable as iron-blast furnace feed, but an equally important objective might have been to furnish an enriched feed for extractive processes for recovering manganese. Flotation concentrates were prepared that contained 11.1 percent manganese, 27.8 percent iron, and 12.2 percent silica, with recoveries of 73 percent of the manganese and 48 percent of the iron. Calculated analysis for a simulated sinter suggests a product containing 16.0 percent manganese, 40.2 percent iron, and 17.6 percent silica, with 31 percent recovery of the dry weight of the ore. The silica in the flotation concentrates was due to quartz and iron silicates. Substantial quantities of the iron silicates (stilpnomelane and smaller amounts of minnesotaite) were present in the slate, and the need to reject these minerals limited the recovery of iron. The sintered flotation concentrate would be usable as ferruginous manganese ore for iron blast-furnace feed because of its favorable manganese silica ratio. The unsintered flotation concentrates, representing 45 percent of the feed weight and containing manganese enriched in the ratio of 1.6 : 1, also could be used as feed to processes for extracting manganese.

Jan 1, 1961

By G. G. Richards

"The slag fuming process has operated for close to 70 years to recover zinc from lead blast furnace slag. The process is normally carried out through the batch treatment of slag by submerged injection of coal and air. Within the bath the coal acts as both a reductant, to reduce Zno to zinc vapour, and a fuel, to maintain bath temperature during the endothermic reduction process. The partition of coal between the slag, combustion and bypass is dependent on factors such as bath depth and injection pressure. The option of continuous fuming has been pursued at the D. Blagoev smelter in Bulgaria using heavy fuel oil as reductant/fuel. There has, however, been little attempt at commercial continuous fuming with coal. This paper presents an analysis of a coal-based continuous slag fuming furnace. The data on the Plovdiv fumer is used as a reference point. Critical design criteria for continuous fuming include the slag feed rate, degree of coal combustion and furnace size. Mathematical modelling analysis suggests that continuous fuming with coal can achieve respectable zinc/coal ratios.IntroductionThe slag fuming process was developed in the early part of this century to recover zinc from non-ferrous smelting slags [1-3]. The majority of installed furnaces were put in place to fume zinc from lead blast furnace slag on a batch basis. In most cases coal is used to effect reduction of zinc oxide, but one furnace uses natural gas [4] and, in a curious twist, a continuous fumer was developed based on fuel oil [5].Given the general nature of blast furnace lead smelting, it is not surprising that slag fuming has remained a batch process, in spite of the advantages that could be gained by moving to continuous fuming. Blast furnaces are usually operated as relatively independent units, and tapping has traditionally been done into ladles. The presence of a crane aisle makes direct coupling of the blast furnace to a slag fumer difficult, especially in a retrofit situation. Furthermore, steady-state blast furnace operation is not easy to achieve. The size of these furnaces and the nature of their feed make it difficult to guarantee constant operation."

Jan 1, 1996

By S. A. Elmquist

Research activities performed by Cleveland-Cliffs Research and Hibbing Taconite plant personnel during the last decade have resulted in the development of an improved autogenous mill grinding circuit. Minus one (1) inch autogenous mill trommel oversize material, con-sidered to be a critical size, is magnetically cobbed and then crushed to 100% minus 3/8" prior to reentry into the feed end of the mill. Autogenous mill capacity increases, up to 25%, have been measured in pilot plant tests using this approach and are also anticipated for the Hibbing Taconite concentrator which presently has reached a 15.5% improvement.

Jan 1, 1994

By M. D. Vilarinho, W. Xuan, F. Vasconcelos Guimarães

"This article summarises the efforts to reduce the energy consumption in grinding in a magnetite ore beneficiation plant. ArcelorMittal Kryvyi Rih ore beneficiation plant produces magnetite concentrate at 65% Fe and 8% SiO2, of which the D80 is approximately 44µm after three stages of grinding with ball mills and magnetic separation. Characterisation and laboratory scale bench tests performed on samples from the beneficiation plant indicate that the industrial flowsheet could be simplified. After the first stage of grinding and magnetic separation, a significant size fraction of the intermediate concentrate is already well liberated. Therefore, it can be directly concentrated in a second stage magnetic separation without further grinding. If a screen of 150µm size can be implemented after the first stage of grinding-magnetic separation, only the oversize of screening, representing 35% of the material, requires secondary and tertiary grinding and magnetic separation. The combined concentrate from the two fractions already meets the industrial specifications. This new route can provide the same final product, with fewer mills; it saves around 30% in energy consumption only in the grinding step without changing the main structures of the processing plant.INTRODUCTION Comminution, including crushing and grinding, is responsible for at least 40% of total energy usage in mining and mineral processing. This paper shows an example to reduce grinding cost, by changing the flow sheet design strategy in ArcelorMittal Kryvyi Rih (AMKR).AMKR is one of the most important mining and steel complexes in Krivoy Rog basin in Ukraine, which has been studied over 100 years because of mining of iron ores. The iron ores are associated to Banded Iron Formation of the Saksagan suite, with an average iron grade of around 36 percent Fe. The total thickness of the Krivoy Rog Supergroup reaches 5-6 km, whereas that of the iron-ore Saksagan Group comprises up to 1400 m. High-grade iron ore extends to more than 2,000 m below present surface"

Jan 1, 2018

By K. Gardner

"Many deposits have a geological model developed early in the project stages, which is not always reviewed or modified as the project matures and production commences. Conversely, ongoing pressures, due to a combination of declining head grades, a push for increased revenue and a multitude of other reasons, leads to processing plant modifications, with the primary aim being to increase production. Often, the ore deposit knowledge is not sustained or developed during the operational stages to the level carried out during the prefeasibility and feasibility study phases. This results in the variability within the deposit being poorly understood and a lack of data and knowledge being available to support (and ideally optimise) operations particularly in an environment where the flow sheet has changed and ore feed requirements are complex.INTRODUCTION This paper presents a case study of ore deposit knowledge development and corresponding process plant operating strategy at the Ladolam Gold Deposit, commonly referred to as the Lihir gold deposit in Papua New Guinea. The gold deposit is located in a tropical equatorial climate on the east side of Lihir Island in the New Ireland Province of Papua New Guinea. Gold mineralisation was discovered in 1982, and was followed by a major exploration campaign by a joint venture between Kennecott (Rio Tinto) and Nuigini Mining Limited. The first drillhole was drilled in 1983, and after an extensive period of drilling consisting of 343 holes for 83,500 m, the final Feasibility Study was completed in 1992. The feasibility study was based on a Reserve base of 37.9 Moz in refractory sulphide ore, with gold extraction proposed to occur via whole ore pressure oxidation due to the refractory nature of the mineralisation. The Lihir Special Mining Lease (SML) was granted in 1995 to a new PNG company – Lihir Gold Limited, and Rio Tinto established Lihir Management Company to develop and operate the mine on behalf of LGL. Mining and processing of the deposit commenced in 1997. Newcrest Mining Limited purchased Lihir Gold Limited in August 2010, and subsequently completed a major plant upgrade including the installation of additional grinding, autoclave and carbon-in-leach capacity. Post commissioning of the upgrade, Newcrest has continued to increase throughput and production. A reinvigorated ore body knowledge development program has significantly assisted this process in identifying operating strategies for the process plant and delivering recovery improvement projects to address declining head grade and recovery due to increased stockpile feed and flotation circuit use."

Jan 1, 2017

By C. A. Sosa-Blanco

Minera Sabinas in central Mexico treats a complex sulfide ore containing galena, chalcopyrite, sphalerite and argentite as valuable minerals, and in some cases important amounts of non-magnetic pyrrhotite. For high pyrrhotite ores, results obtained by selective flotation showed good grades and recoveries in the lead-copper circuit, but a commercial zinc concentrate could not be produced due to the high floatability of pyrrhotite. When this is the case, an alternative method was proposed to produce a bulk concentrate by flotation which is then leached to recover copper and zinc from liquor in a SX-EW circuit. Leaching was conducted in an autoclave in order to study the effect of temperature and pressure, but bioleaching tests were also conducted. Leaching lead-silver residues were treated by cyanidation in order to recover silver values. The results obtained by selective flotation and by bulk flotation with leaching of the bulk concentrate showed that the latter alternative increased zinc recovery by 30 % with similar recoveries of the other metals.

Jan 1, 2007

By Peter Kreisman

This paper and four other papers on the MEQB Regional Copper-Nickel Study, presented today at the 1980 SME-AIME Fall Meeting are based on four years of extensive inter-disciplinary research and impact analysis supported largely by the Legislative Commission on Minnesota Resources. This study produced over 180 technical reports, the results of which are summarized in an Executive Summary available from the Minnesota State Planning Agency, St. Paul, Minnesota. Those wishing to examine the many scientific, engineering, environmental, social or economic issues addressed by the study are encouraged to obtain a copy of the executive summary. The executive summary includes a bibliography of all technical reports produced by the Regional Copper-Nickel Study and describes the five-volume, 26 chapter Regional Copper-Nickel Study Report.

Jan 1, 1980

By Stephen P. Holt, John A. Bowsher

A preliminary examination of the mica deposits in Culberson and Hudspeth Counties , Texas , was made in September 1942 by A. B. Needham, district engineer of the Bureau of Mines . The deposits were reexamined by R. C. Redfield , associate mining engineer of the Bureau in November 1942, at which time he was accompanied by Philip S. Hoyt , geologist , and J. C. Moore , foreman of the Texas Mica and Feldspar Company . Exploratory work was started on February 16 , 1944 , and terminated on April 30 , 1944. A crew of 4 to 10 men was employed under the supervision of the authors .

Feb 1, 1947