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By K. S. Fraser, H. J. H. Pieterse

"A pressure oxidation plant to pretreat refractory sulphide ore at the Goldstrike plant in Nevada, was commissioned in early February 1990. The plant, with a nominal capacity of 1363 t/d, was commissioned over a one-week period and achieved design capacity and availability within ten weeks an unusually short period for a plant of this complexity.The pressure oxidation process is conducted under elevated temperature and pressure conditions, and in a corrosive environment. These conditions demand that the process and engineering design of the plant be carried out in a much more sophisticated and detailed manner than is normally required in a conventional gold plant.This paper provides an account of the process and engineering design of the Goldstrike pressure oxidation plant with special emphasis on the design of the vessels, slurry pumps, piping, valves, and control systems.BackgroundLocationThe Goldstrike property is located in Eureka County in northcentral Nevada on the Carlin geological trend. The mine is approximately 40 km northwest of the town of Carlin which is situated on Interstate Highway 80. The minesite is located 1700 m above sea level in the Tuscarora Mountains.Initial Goldstrike OperationsShallow, low-grade mineralization was discovered by diamond drilling in the Goldstrike area in the 1960s. A heap leach operation was established in the mid-I970s by the previous owners. The Post deposit was discovered in 1982. Gold production by mining and heap leaching reached I t/a (40 000 oz/a). American Barrick purchased the property in 1987 and started an aggressive program of deep drilling. This led to the discovery of the Betze and Deep Post deposits and established sufficient reserves to justify an initial 4100 t/d mill and a 5 500 000 t/a heap leach operation"

Jan 1, 1992

By W. J. Sharwood

The assay or estimation of the total gold content of an ore presents little difficulty, when circumstances permit of securing a thoroughly representative sample. The general practice has been fairly standardized, and such errors as occur in the results obtained are for the most part attributable to the methods of sampling rather than to those of assaying, while duplicate assays made by skilled workers usually agree closely. When it is a question of estimating the "free" or amalgamable gold of an ore, as opposed to the total gold, there is but little uniformity in the methods employed by different workers, and often equally little in the results obtained when duplicate determinations are attempted, no satisfactory standard method having been generally adopted. The methods in use are of two general types—"panning" and "amalgamation." Panning For the estimation of the value of the visible gold by panning or washing, a weighed sample may be taken, ranging from an ounce to several pounds; a kilogram or 2 lb. is a favorite quantity for laboratory tests, while a prospector often uses a full miners' pan, say nearly 20 lb. or 10 kg. In the case of rich quartz, ground to about 30 or 40 mesh, a Mexican horn is convenient, using 1 to 8 02;. of material, and a 6-in. frying pan is favored by many prospectors in dry districts. Bateas, plaques, and dishes of various patterns are used for the same purpose. With these, fairly close results may be obtained by using for ocular comparison standards consisting of weighed amounts of gold of similar character as regards purity, fineness of division, and the general shape of the particles, which standards are preferably obtained from gold of the same locality, or the same deposit, as the ore or gravel being tested. Men working in one particular district sometimes become surprisingly expert in thus judging the value of an ore by the eye, the weights of both ore and gold being guessed. On the other hand, many have been led to overestimate the value of ore and gravel from unfamiliar localities, where the gold has been very finely divided or of unusually high silver content, and a novice may

Jan 1, 1916

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By O. D. Sherby, J. L. Lytton, C. R. Barrett

Randomly oriented polycryslalline copper of 99.995 pcl was tested in tension at temperatures of 626o, 496o, and 406o. The gvain-size mnge investigated was from 0.03 to 0.7 mm. Grain sizes were produced by two techniques: 1) varying the amount of prior cold work and the amealing time at constant annealing temperature so as to obtain various vecrystallized grain sizes with minimum grain growth, and 4 holding the prior cold work constant while varying the annealing time or temperature so as to obtain varying grain sizes by grain growth. Polycrystalline samples (grain size of 0.03 mm) with a strong. [001](100) texture were also studied. The relationship between the steady-state creep rate, Es, and pain size was found to be the same indepentent of the technique used to prepare the various grain sizes. For specimens with grain diameters above about 0.1 mm ea zuas esse)ztiallj~ independent of grain size, while for smaller grain diameters Cs increased slightly with decreasing grain size. Strongly textured polycrystalline copper, which contained low-angle grain boundaries, exhibited steady-state creep rates that were slightly lower than those observed in randomly oriented copper, of the same gain size. The results are explained by conside?%ng- the contribution of grain boundary shearing to the total strain and the effect of grain size on the resulting creep substructure. EXPERIMENTAL observations concerning the influence of grain size on steady-state creep rates are varied and oftentimes conflicting. Numerous early investigators found that an increase in the grain size resulted in a decrease in the high-temperature steady-state creep rate, i,.'-" Other workers have foundthat ef decreases with increasing grain size up to some optimum grain size and then increases with a further increase in grain size.'- '' It has been suggested that es, decreases with an increase in grain size only because the range of grain sizes studied normally lies to the left of the grain size for optimum creep resistance. However, sherbyl2 and Feltham and his co-worker13, 14 have found that even for small grain sizes 2, does not always decrease with an increase in grain diameter but may be proportional to the square of the grain diameter. A possible explanation for the different creep rate-grain size relationships which have been observed is that in most experimental studies different grain sizes have been prepared by annealing at various temperatures. For metals in which the impurity concentration is high (greater than about 0.1 at. pct) this type of

Jan 1, 1968

By O. G. Olvera, D. G. Dixon, E. Asselin, L. S. Quiroz

Dissolution rates of chalcopyrite (CuFeS2) were calculated using electrochemical impedance spectroscopy (EIS). The calculated dissolution rate decreased by a factor of 3.2 after the electrode was passivated. The results presented in this work indicate that the reduction of Fe3+ was the rate-controlling step for fresh CuFeS2 electrodes while the transport of cations within the passive film limited the dissolution rate for the passivated electrode. The transfer of cations at the electrode/electrolyte and passive film/electrolyte interfaces did not affect the dissolution rates of the fresh and passivated CuFeS2 electrodes, respectively. The catalytic effect of pyrite (FeS2) on the dissolution of CuFeS2 continued after the passivating treatment. The contact of fresh and passivated CuFeS2 electrodes with different FeS2 particles resulted in increased dissolution rates, due to an increase in the reduction rate of Fe3+.

Jan 1, 2014

By Masaaki Naito

The Japanese steel industry has a long history of introducing new and innovative technologies in the field of ironmaking. The new technologies introduced during the past ten years include technologies to use cheaper and lower-grade raw materials and fuels, measures to prolong the service life of blast furnaces and coke ovens, promotion of energy saving, use of wastes and solutions to environmental problems. This report outlines the condition of production and technological trends and technical development themes in ironmaking technologies.

Jan 1, 2006

By J. A. L. Napier

The application of excess shear stress as an aid to the design of mine layout is reviewed, and the critical parameters required to describe slip on a plane of weakness are noted. A design procedure that is required to model the cyclic build-up of excess shear stress is described. A number of simple examples of mine layouts on two reef planes intersecting a fault plane are then considered to illustrate the application of the design concept and to highlight some of the pitfals that may be encountered

Jan 1, 1987

By A. J. Griest, P. D. Frost, H. A. Robinson

The flow characteristics and uniform elongation of alpha-beta titanium alloys in the solution-treated condition were shown to be markedly affected by the solution temperature. Two classes of alpha-beta alloys were distinguished, based on the effect of solution temperature on the uniform elongation. Martensitic alpha-beta alloys, which are relatively weakly beta-stabilized, showed decreasing values of uniform elongation as the solution temperature was initially increased. However, after solution-treatment at still higher temperatures in the alpha-beta field, the uniform elongation of these alloys was again high, and often attained its maximum value. Nonmartensitic alpha-beta alloys, which are more heavily beta-stabilized, showed a continuous decrease of uniform elongation with increasing solution temperature. The behavior was rationalized in terms of the alpha-beta ratio and the alloy content of the beta existing at the solution temperature. The data indicate that in commercial practice close temperature control may be required to get maximum uniform elongotion, and hence, maximum potential formability, in solution-treated sheet. Bend-test data also were obtained for the solution-treated conditions. A MOST important objective of current titanium research is the development of age-hardenable alloys with superior formability in the solution -treated condition. Alpha-beta titanium alloys have been shown to be amenable to high-strength aging treatments after solution-treatment in the alpha-beta field.' However, heat-treatment data on the properties of solution-treated sheet, such as uniform elongation, minimum bend radius, and flow properties, which have a bearing on the potential formability, have not been widely available. This paper presents only the most significant results of a research program conducted at Battelle for the Air Force in an effort to develop sheet alloys that can be formed readily as solution-treated and subsequently aged to very high strengths. Data for 10 compositions representative of the more than 25 alloys studied in this program are presented. A primary objective of the paper is to demonstrate the pronounced dependence of the flow characteristics and uniform elongation on solution-treatment temperature. This effect is believed to be vitally important in the commercial heat-treatment of age-hardenable titanium sheet alloys. Experimental Procedures Melting and Fabrication—The nominal and actual compositions, beta transi temperatures, and rolling temperatures of 10 of the 30 alloys studied are presented in Table I. Stock :€or two of the alloys, E (Ti- 8Mn) and J (Ti-6A1-4V) was obtained from commercial heats. The remaining heats, except D and I, were double melted as 10-lb heats in a consumable-electrode furnace. Melting stock was 120 Brine11 hardness sponge. The oxygen and hydrogen analyses spotted throughout the heats indicate the range of interstitial levels resulting from the melting practice used. Heats D and I were melted as 20-lb ingots with the same grade of sponge. These ingots were cut in half and one portion used in the present evaluation. The ingots, which were about 4 in. diam and 4 in. in height, were upset forged and drawn at 1750°F to sheet bar 5/8 in. thick by 2 1/2 in. wide in such a manner that the top of the ingot corresponded to the top of the sheet bar. Slabs of the sheet bar, 5 1/2 in. in length were scalped to 1/2 in. thickness and rolled at the temperatures indicated in Table I to approximately 0.055 in. sheet. Elongation during rolling took place at 90" to the original forged length. The fabrication procedure used is described in some detail since the experimental heats showed considerable directionality in mechanical properties. The rolling procedures for the commercial heats were unknown, but were such as to produce essentially isotropic properties. All heats, as rolled, showed microstructures typical of alpha-beta alloys worked in the alpha-beta field; that is, the structures consisted of equiaxed alpha particles in a beta matrix. Heat Treatment— The effects of solution and aging heat-treatments on the properties of the alloys were evaluated uSing longitudinal and transverse tensile and bend coupons sheared from the sheet. Two or three solution-treated conditions were evaluated for each alloy. In addition, the alloys were also tested after annealing at 1200°F for 4 hr and furnace cooling. The solution-treated and annealed conditions were those in which the alloys might be expected to

Jan 1, 1959

By R Morrison, P Walters, B Adair, V Jokovic

‘Grade Engineering<sub>®</sub> is an integrated approach to coarse rejection that matches a suite of separation technologies to ore specific characteristics and compares the net value of rejecting low value components in current feed streams to existing mine plans as part of a system-view’ (Walters, 2017). Two of these technologies are directly recoverable through screening stages: Natural Deportment and Induced Deportment. The potential for taking advantage of “Natural Deportment” by size following blasting and crushing has been demonstrated for a range of ore types. Simple and low cost screening can in some cases generate a fine product with improved grade and a lower grade oversize which can be directed to leach or to waste. Where an ore body consists of many different ore types, the traditional approach of blending to metal grade will often add to complexity of treatment, which will reduce the effect of “natural deportment” by size.CITATION: Jokovic, V, Walters, P, Adair, B and Morrison, R, 2018. Applying enhanced grade engineering to complex ore bodies, in Proceedings Complex Orebodies Conference 2018, pp 21–23 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018

By R. W. O. Kersten

The main objective of The South African Institute of Mining and Metalurgy is to identify the needs of the mineral and metals sector of the South African economy, and to initiate the means by which the requirements for technology and scientific knowledge can be satisfied. To achieve this, the Institute organizes schools to promote technology transfer and to satisfy the needs of its members for continuing education by the exchange of technical information, the communication of ideas, and the identification of technological problem areas.

Jan 1, 1986

By R. M. Tyson

Sulfides present in the troctolitic and footwall rocks are placed into one of three major zones (hornfels, cloud, and basal) on the basis of their stratigraphic position and mineralogy. Hornfels sulfides, enriched in Fe, represent remobilized sedimentary sulfide. Sulfides from the Cloud Zone are probably magmatic and are enriched in PGE, Cu, Ni, and Ag. The basal sulfides, composing most of the ore, resulted from contamination of the magma by remobilized sedimentary sulfide. The basal sulfides exhibit a differentiation trend in their mineralogy and chemistry. PGE, Cu, and Ag are enriched with increasing differentiation which is attributed to fractionation within an immiscible sulfide liquid.

Jan 1, 1986

By K. Becerra, D. G. Almond, T. D. Maru

"Nkomati Nickel Mine in South Africa processes one of the world’s hardest ores in a 375k TPM autogenous grinding circuit. The plant employs an unconventional processing circuit, comprising a primary autogenous mill (AG or FAG) and pebble crushing circuit followed by a secondary hybrid pebble/ball mill and classification cyclone cluster.Advanced Process Control (APC) or Expert Control is now generally accepted as an advantage in most modern high capacity concentrator grinding circuits. This paper reviews the installation of model predictive control at Nkomati and examines the results and benefits of installation in an autogenous mill during plant ramp-up. The paper also demonstrates how predefining process control objectives together with good plant and supplier cooperation can achieve a robust and reliable APC implementation on a very short site program.The results conclusively demonstrate operational and economic benefits during circuit operation with the advanced process control system ON. Production performance and circuit stability were significantly better than during manual control."

Jan 1, 2012

By F. E. Wagner, L. J. Cabri, Y. G. Safonov, N. S. Bortnikov, A. D. Genkin, J. Friedl, G. N. Gamyanin

"A multidisciplinary investigation (including INAA, SIMS, EPMA, and Mossbauer Spectroscopy) was used to characterize gold in four of the largest Siberian mesothermal sulfide gold deposits. These are the Olympiada (""'750 metric tons gold) and Veduga (estimated at 75-100 metric tons gold) deposits in the Enisei Mountain Range and the Nezhdaninskoye (z500 metric tons gold) and Sentachan deposits in the Verchoyansk-Kolyma area. It was found that some of the mineralization is refractory in all four deposits, largely due to the occurrence of ""invisible"" gold in arsenopyrite as well as within pyrite, but with a lesser concentration. The gold concentration and its distribution varies considerably both within and between arsenopyrite grains, with implications for processing options such as heap leaching. The ore most likely to be amenable to heap leaching is the oxidized ore of the #4 orebody of the Olympiada deposit, but further quantitative mineralogical investigations are recommended for this orebody's sulfide ore and for the ores of the Nezhdaninskoye deposit. Following a full mineralogical characterization, it is possible that economic viability will depend on a combination of pre-concentration with or without pre-treatment, followed by heap leaching of some tailings products.Acknowledgements· The Russian authors performed this study under an ""Eco res"" Project which was financially supported by the Ministry of Science and New Technology of the Russian Federation. C.J. Stanley wishes to thank John Spratt and Terry Williams for help with EPMA. F.E. Wagner acknowledges the Deutsche Forschungsgemeinschaft for a grant supporting the Mossbauer studies. L.J. Cabri is grateful to the following: M. Beaulne, D. Carson and J.H.G. Laflamme for technical support, K.E. Haque for technical comments, and D.M. Hausen and W. Petruk for editorial suggestions."

Jan 1, 1997

By N. Razavinia

Most mining and metallurgical industries produce relatively large quantities of low grade waste heat which they vent, for the most part, to the environment. To date, little is being done to recover some of this heat and transform it from low grade to a high grade energy form such as electricity. A few decades back this was also true of paper, plastics and metals. Recycling was not a prominent activity. Today, we all know that it is and that recycling is a must for sustainable development. To this end, we feel that we are now at that point in time when the recycling of waste heat to electricity must be given serious consideration. By recycling waste heat one can achieve at least one if not two of the following benefits: 1) reduction in the carbon footprint ? by improving the overall energy efficiency of a process, and 2) production of electricity at an economically viable cost. We have developed a process that captures, concentrates and converts (3 C?s) waste heat to electricity. The process comprises 3 distinct interacting units. The first unit captures and concentrates waste heat using heat pipe technology. The second unit converts some of the concentrated heat to mechanical work which can then be readily converted to electricity. This unit is based on the classical piston/cylinder engine arrangement that is widely used by consumers to power systems with compressed air, for example. The third unit of the process dissipates the heat that the engine cannot convert to high grade energy. It also uses heat pipe technology to achieve this. The combination of the 3 units forms a unique system for recycling waste heat. This paper describes the system and it presents some results that show how the system can be made to be economically viable.

Aug 1, 2013

By R. W. Watson

This Bureau of Mines report summarizes some recent results on the sensitivity of various experimental explosive systems using ammonium nitrate (AN) as the oxidizing agent. The sensitivity survey included drop-weight and friction testing and sensitivity to initiation by projectile impact and a No.8 blasting cap. It was found that certain types of AN prills can result in cap-sensitive mixtures when combined with No.2 diesel fuel (Fa); prilled AN mixed with nitromethane also result in cap-sensitive mixtures. Mixtures of AN prills and neat 1-and 2-nitropropane were not cap-sensitive, and the use of powdered or grained aluminum in AN-FO did not adversely affect the sensitivity except possibly to increase its sensitivity to friction.

Jan 1, 1974

By Alexandr I. Volkhin

It is practically useful to separate copper from nickel at the fire refining stage of blister copper by transferring nickel in electrochemical inert silicides which are being concentrated further into the tank-house slimes. The treatment of copper melt by silicon-containing reagents is resulted in lowering the interfacial tension of the melt, its deoxidizing, selective siliconizing of in-melt nickel and forming a by-product with fluxing properties. Si-containing reagents have positive influence on results of fire and electrolytic refining, and further pulverizing of copper in plant conditions.

Jan 1, 2006

By Thomas H. Sheldon

In the zeal for opening up new ore-bodies, or for extracting the ore from attractive bodies already opened up, we very often lose sight of the fact, that, after all, the operation of a mine is a business proposition, pure and simple, and, for the best working-results, should be treated upon a strict business basis. Of course, in every mine of consequence, a record is kept of expenditures and receipts, and such glittering generalities as " gross receipts," " net receipts," " mining expenses," and a per cent. profit," can be told to the cent; but does this record show economy of management, as compared either with the same record of other months, or with the record of other mines of the same class ? Moreover, if such a record shows that the cost of mining is high, does it in any way enable the manager to put his finger on the leakage ? Does it necessarily follow that a mine which makes a profit of, say, 40 per cent. on 25-dollar ore, is doing less economical mining than one which saves 60 per cent. on 80-dollar ore ? Of course the figures in the latter case look the more attractive; yet when it comes to the point of saving everything which can be saved, and of cutting down expenses to the lowest possible cost of operation, the former mine is doubtless on the firmer and more economical financial basis. But as to the relative merits of the system of mining in the two cases, nothing could be decided without a basis of detailed comparison ; if one system is more economical than the other, why is it so, and wherein does the advantage lie ? This can be shown only by keeping accurately the cost of each mining operation. And no matter how dissimilar two mines may be in character and operation, yet there are a few general heads common to all mining operations. In the first place, it is necessary to break the ore from the solid ground;

Jan 1, 1907

By Alan Lawley

AS part of a programme of experimental work on single crystals of rhenium and ruthenium, it was found necessary to have the crystallographic angles for these elements. The angles given in Table I were determined to the nearest minute of arc, by Univac Computer. The angle $ between (HKIL) and (hkil) is given by the formula: The angles given for rhenium should be particularly useful in view of the increasing interest in this element. For completeness, angular values are also given for cadmium zinc and magnesium; these form a more complete set of values than those given by Salkovitz.' The assistance of Professor J. Hobstetter and Mr. J. Crozier in the programming of this data is acknowledged. The work is sponsored by the Office of Naval Research under contract No. NONR 551 (19).

Jan 1, 1961

By Brij M. Moudgil

Separation of oil shale from waste material or into high grade and low grade fractions can have a major influence on the economics of the surface (ex-situ) retorting process. Different components of shale rock vary significantly in physical and surface chemical properties such as specific gravity, color, water wettability etc. Various techniques which utilize the difference in one or more of the above propartis have been proposed to preconcentrate the oil shale before retorting. A discussion of the techniques proposed such as gravity separation, sorting and froth flotation is presented.

Jan 1, 1982

By A. D. Hoagland

The Flat Gap mine is located on Copper Ridge in the Appalachian Valley of East Tennessee. Large zinc orebodies occupy zones of solution and collapse breccias in Lower Ordovician Kingsport limestone and dolomite. The principal ore mineral is sphalerite and the principal gangue is dolomite. The ore-bearing horizons are considerably more than 95% covered by thick residual clay. Geochemical prospecting under these conditions is an effective exploration technique which played an important role in the discovery and exploration of the Flat Gap mine. The Flat Gap mine, an important zinc deposit of the Appalachian type, is located in the Copper Ridge Belt in the Valley of Tennessee approximately 30 miles northeast of the Mascot-Jefferson City zinc district (Fig. 1). The ore is found in large stratigraphically linked deposits in solution and collapse breccias of the Kingsport, Longview and, to a minor extent, Mascot formations of Lower Ordovician age. The principal ore mineral is sphalerite and the principal gangue minerals are dolcrmite and accessory pyrite. There is local minor galena and very rare barite and fluorite. The ore-bearing horizons are covered by about 20 to 100 ft or more of residual clay except where active streams have exposed bedrock. Considerably less than 5% of the favorable host rocks are exposed in outcrop, and this nearly complete soil cover creates a serious obstacle to exploration. Comprehensive surveys of the base metal content of soils furnish a useful substitute for outcrop geology, and under favorable conditions, this technique is a very valuable prospecting tool. The analysis is carried out by a titrametric dithizone method modified after the procedure described by E. B. Sandell.' This method detects

Jan 1, 1962