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By J. A. Rowland, C. W. Funk

THIS investigation is a part of the United States Bureau of Mines work in conserving the Nation's resources. The isothermal sections presented were developed as a guide to a comprehensive investigation of the properties and fabricating characteristics of copper-base alloys containing manganese and tin. These alloys are being investigated as possible replacements for the commercial bronzes containing substantial quantities of tin. Development of the isothermal sections has, therefore, been limited to the area between 0 to 20 pct Mn and 0 to 25 pct Sn. Since Heusler1 reported the ferromagnetic properties of the Cu-Mn-Sn alloys in 1903, the system has been the subject of several investigations. These studies were largely concerned with correlation of magnetic properties and crystal structure.2-6 Vero,' however, established the solidus surface of the system, between 20 pct Mn and 40 pct Sn, by thermal and microscopic methods. The a, ß, and y solid solutions were also observed by Vero and reported to extend into the interior of the ternary system. For the present study, the annotated diagram of the Cu-Sn system prepared by Raynor' and the Cu-Mn diagram by Dean and coworkers9 were used as the binary borders of the ternary system. Procedure The metals used in this investigation were electrolytic manganese, washed cathode copper, and three-star tin. The manganese, produced at the Boulder City pilot plant of the Bureau of Mines, had a purity of 99.9 pct; the copper contained less than 0.004 pct Bi and 0.001 pct S; the tin contained 0.070 pct Pb, 0.070 pct Sb, 0.050 pct Fe, and 0.065 pct Si. Heats of 1 lb were melted in alundum crucibles with a 3-kw induction furnace. The heats were chill-cast in iron or copper molds to form 6-in. ingots 3/4 in. in diameter. Molds were washed with graphite or zirconia, depending on the manganese content, and preheated to 150°C. The maximum impurity content of any alloy was 0.03 pct Fe, 0.02 pct Si, and 0.02 pct Al; in most cases, they contained less than half this quantity of any of these impurities. Composition of the alloys and their treatment prior to homogenizing are shown in Fig. 1. Designations adjacent to each composition refer to the heat numbers. With a few exceptions, ingots containing 20 pct Sn or less were hot-swaged. These ingots were swaged, after a 24-hr preheat at 650°C, to obtain a total reduction of 70 pct in cross section. Intermittent reheating was necessary, and reductions were limited to either 0.025 or 0.050 in. in diameter per pass. Selection of a 200-hr homogenizing treatment was based on extensive experiments which indicated that virtual equilibrium was reached after 150 hr. Structures were retained by quenching in water. Grain sizes in the specimens homogenized at 350" and 450°C were generally small. These homogenizing treatments were preceded by a 50-hr anneal at 650°C to facilitate phase identification by increasing the grain size in these specimens. The specimens were furnace-cooled from this temperature to the homogenizing temperature, and the heat treatment was continued for the standard 200-hr period. The homogenized samples were bisected to provide an internal section for metallographic examination. Optimum definition of the microstructure was obtained by successive polishing and etching. Quarter-strength ASTM copper reagent No. 13 gave the most satisfactory results.10 Filings from interior sections of the samples were used for diffraction studies. These filings were annealed in evacuated glass tubes at homogenizing temperatures for periods exceeding 50 hr and quenched in water. During this treatment the manganese content was reduced by significant amounts, probably by vaporization. This change in composition was considered in applying the diffraction data, which were used only as a means of identifying phases where the data were consistent with the metallographic evidence. Phase Identification The a solid solution is readily distinguished by the copper-colored, twinned, polyhedral grains, as shown in Fig. 2, and a face-centered cubic X-ray pattern. Tukon indentations, using a 25-g load, indicate a Knoop hardness of 155 for a quenched from 650°C. Using white light developed by a Wratten 78 A filter, the ß grains in etched specimens have a dark brown tint in contrast to the lighter copper-colored grains of the a phase. Knoop hardness values of 284 were obtained for the ß structure in specimens quenched from 650°C. The ß structure was also distinguished from the a phase by its body-centered cubic lattice. The acicular structure appearing in Fig. 3 was evident in several alloys of higher tin content after quenching from either 750' or 650°C. Since a trans-

Jan 1, 1954

By Bawden WF, Paynter JT

This investigation involves laboratory testing and stability analysis of paste fill at the Golden Giant Mine. Site-cast samples are tested in unconfined, confined, and consolidated-undrained triaxial compression. Pore pressures during shear and consolidation characteristics are measured to determine both drained and undrained shear strength parameters and assess the potential for undrained loading. Liquefaction potential is also considered. A limit equilibrium solution is used to obtain factors of safety against failure of the fill during mining of adjacent stopcs

Jan 1, 1998

By Harold N. Underhill

Larger equipment, more attention to maintenance, new types of bits, radio communication, and advances in jet piercing highlight an active year.

Jan 3, 1951

A major problem in open pit mine planning is to determine the optimum pit limit of the mine. The optimum pit limit is the one with maximum discounted net profit and satisfies technical conditions. A number of optimisation techniques have been proposed for determination of pit limits. Among these techniques, those based on graph theory are suitable to solve the three-dimensional (3D) problem. In this paper, two maximum flow methods named, Dinic's algorithm and Push-Relabel algorithm have been used to find the optimum pit limit of an iron ore mine. This method has been implemented in Gol-e-Gohar iron mine in south east of Iran in order to find the final pit limit. Issues, like properties of the mine and orebody, are studied. These properties include wall slope, sulfur and phosphor grade restrictions. The result shows that these properties have significant effect on the determined pit limit.

Dec 6, 2010

By Fidelis T. Suorineni

HydraboltTM is a Swellex-type dowel support element developed and manufactured by New Concept Mining in South Africa. Like Swellex, the HydraboltTM is a folded tube which can be inserted into a larger diameter hole without pushing. It is then activated by injection of high pressure water which inflates the folded tube to intimate contact with the walls of the borehole at a set pressure, normally 25 MPa. Unlike Swellex however, the set water pressure at the end of the inflation is maintained throughout the life of the dowel. An integral non-return valve makes this possible. The HydraboltTM is being investigated as a possible stress change measuring cell. This paper presents the results of closed form solutions, numerical modeling and laboratory test to date. Preliminary results indicate that the HydraboltTM can be used as a stress change monitoring cell in addition to its conventional role as a rock mass reinforcing element. If successful, the HydraboltTM will supplement microseismic monitoring systems in rockburst-prone ground conditions where they will give quantitative stress change data for both mine design model calibrations and means for proactive safety measures.

Oct 1, 2010

The drilling and evaluation of bore cores provides the best available statistical basis for characterizing a coal prospect. Important, long-term decisions by management frequently must be based on the results. The majority of coal mined presently is used for steam raising or the production of metallurgical coke. Common bore core tests and analyses appropriate to this usage can be cate- gorized as non-destructive, chemical, physical, petrographic, caking/plastometric and propert- ies of non-coal material. The major tests within each category are described. Certain specific information from the bore cores is required by the coal geologist, mining engineer, coal preparation engineer and the user, either individually or as a group. The optimal bore core evaluation test pro- gramme is the one which satisfies the needs of all the professional disciplines with the grtatest speed and at minimal cost. The principal steps in such a programme are listed. There is an urgent need for a bold new Australian standard to provide the necessary guidelines to an industry which at present lacks uniformity in its approach to the task.

Jan 1, 1977

By K. D’Elia, L. D. Ogrel, O. A. Malyshev, T. I. Yushina, I. M. Petrov

"The purpose of this research is to develop enhancements in the mineral processing by flotation methods of nonferrous ores. Currently, 180-190 million tons of ore are processed by flotation in Russia. The annual processing volume of copper-nickel and copper-zinc ores exceeds 20 million tons. In the total volume of processed ores by flotation, non-ferrous ores account for roughly 35% and ores of precious metals for another 15%. The total demand of flotation reagents in Russia in 2016 was estimated at 170,000 tons. Sulfur-based collectors used in the flotation of non-ferrous sulfide ores (typically, xanthates) do not always offer sufficiently high recovery of valuable components.It has been found that reagents like DMIPEC and DK-80 improve the recovery of metals into concentrates. To identify the reasons of that higher recovery, quantum chemical calculations of the mechanism and interaction forces of DMIPEC with non-ferrous sulfides were undertaken by determining the stabilization energy of the resulting chemical complexes. It has been shown that, in the interaction of DMIPEC and DK-80 with the surfaces of sulfide minerals, ethylene and acetylene bonds act as the reaction centers, and the formation of molecular p-complexes is the result of metal-cation interactions with the surface and the flotation reagents.In the flotation of non-ferrous ores, DMIPEC and DK-80 act as selective supplementary collectors with foaming properties, and at a flow rate of 3-35 g/ton, they reduce the flotation time, and increase copper recovery by up to 3.5%, zinc recovery by up to 12%, gold recovery by up to 8-10%, allowing to increase the recovery of valuable components by 1.5-12.0%, basically this kind of reagents nearly double the flotation rate, improve the grade of the concentrate, improve the stability of the flotation process parameters, and reduce the environmental impact of flotation plants."

Jan 1, 2018

By Darren Medd

Kumtor Gold Company (KGC) currently operates the 500,000 tonnes per day Kumtor open pit gold mine located in the Issyk-Kul region of the Kyrgyz Republic at an elevation of more than 4,000 m. To supplement the open-pit mill feed with high grade underground ore from outside the current pit limits, KGC started pre-production design and development of the underground mine for the SB ore zone in late 2006. The initial phase of pre-production development consists of exploration and haulage declines which, upon completion, will provide ventilation and access to the underground mine. Design and construction of these declines presented many challenges to the project team considering the mines unique location at high altitude, the limited experience with Western underground mining techniques in the Kyrgyz Republic, and the impact of the nearby Davydov glacier and resultant ground conditions. Completion of the exploration and haulage declines was undertaken jointly by KGC and mine contractor, J.S. Redpath Limited (Redpath). The initial phase of development including several ground falls provided detailed information on geotechnical ground types and conditions that were useful in determining the overall mine design and methodology for production. This paper covers the various aspects of the design and construction process of developing the declines under the glacier, minimizing the likelihood of further ground falls, and the journey of both KGC and Redpath which led to the successful breakthrough between the declines on 01 July 2012.

Aug 1, 2013

By Christopher Mark, Anthony T. Iannacchione

Falls of ground continue to be one of the most serious causes of injury to U.S. miners. Of the 256 fatal injuries that occurred in mining between 1996 and 1998, 52 (20%) were caused by falls of ground (Table I). Falls of ground affect some sectors of the mining industry more severely than others. For instance, nearly 40% of the 98 coal mine fatalities between 1996 and 1998 were caused by falls of ground. Underground miners are at much greater risk than surface miners. Nearly half (45 out of 101) of underground mine fatalities were attributed to roof, rib and face falls, while less than 5% of the 155 surface fatalities were caused by falls of highwalls or slopes. Sources of Data All of the injury data examined in this study were derived from MSHA’s Fatal Investigation Reports (www.msha.gov)and the MSHA accident database. Because falls of ground often result in serious injury, MSHA Fatal Alert Bulletins and Fatal Investigation Reports provide a useful snapshot of ground control issues in the mining industry. MSHA requires that mines file a report on every reportable accident that occurs, containing information on the accident’s location, severity, classification, activity, and nature of injury, etc. A short narrative is generally included as well. Accident reports can be searched by many of the above fields. From 1996 to 1998, miners suffered a total of 55,096 injuries, which ranged in severity from death (degree 1) to injuries with no days away from work nor restricted duty (degree 6). Six percent of the total injuries were from falls of ground, including machine accidents where caving rock was coded as the source. As the data in Table II indicate, 98% of all non-fatal fall of ground injuries occurred in underground mines, with underground coal mines accounting for 83% of the total. Table II also shows the distribution non-fatal fall of ground injuries by severity and commodity. The injuries are classified into Lost Time Injuries that resulted in permanent disability (degree 2) or days off work (degrees 3-4), and Injuries Without Lost Time that resulted in no more than restricted duty (degree 5-6). Overall, groundfall injuries appear to be more serious than other types of mining injuries. 65% of all ground fall injuries resulted in lost time compared to 54% of all types of mining injuries.

By G. L. Tufford, R. L. Bleifuss

Low-grade manganiferous iron ores of Minnesota's Cuyuna Range, in general, do not respond to conventional mineral dressing techniques because of their fine-grained texture. Reducing these ores to magnetite with subsequent magnetic separation and flotation have not given satisfactory results because, in many instances, concentrate grade and manganese recovery are low. Metallization of iron and concurrent reduction of higher manganese oxides to MnO, as in the R-N process, followed by recovery of iron by magnetic separation and manganese by leaching, seems to be a technically feasible process. High-temperature x-ray diffraction studies show that solid-state reactions between iron and manganese, with the formation of intermediate iron-manganese oxides, occur during reduction and affect recovery of both iron and manganese from the final product. Solid state reactions also occur between FeO, MnO, and SiO, at temperatures as low as 900°C. The kinetics of these reactions are such that they become an important factor in the treatment of ores by the R-N process. These investigations suggest a temperature and gas composition cycle that would maximize both iron and manganese recovery. During the past ten years, iron ore shipments from the Cuyuna Range of Minnesota have decreased sharp1y. Because of adverse grade, structure, and cost, many Cuyuna ores are no longer competitive with high-grade foreign ores and domestic pellets. The low-grade manganiferous iron ores from the Cuyuna Range, in general, do not respond to conventional mineral dressing techniques because of their fine-grained texture. Reducing these ores to magnetite for subsequent magnetic separation and flotation has not given satisfactory results because concentrate grade and manganese recovery are often low. However, a technically feasible process appears to be metallization of the iron and concurrent reduction of manganese oxides to MnO with recovery of metallic iron by magnetic separation and extraction of manganese from the tailings by leaching. In 1962 the R-N Corp., with a grant from the Area Redevelopment Administration, conducted an amenability study of the metallization of manganiferous iron ore from the Cuyuna Range. The results of this preliminary investigation were sufficient to interest the Economic Development Administration in sponsoring further studies of the process with emphasis on secondary recovery of manganese from the nonmagnetic tailings. Inasmuch as the metallization process has considerable influence on the final form of the manganese - its texture and distribution between the magnetic concentrate and nonmagnetic tailings-the investigation included a study of metallization of the ore as well. Therefore, the investigation conducted at the Mines Experiment Station covered three interrelated phases: (1) laboratory R-N roasting studies and magnetic separation tests, (2) laboratory chemical extraction experiments of the nonmagnetic tailings, and (3) a mineralogical study of the crude ores and their various roast products. High-temperature x-ray diffraction investigation of the Fe-Mn-Si0,-0, system is the main topic of this paper and an extension of the mineralogical work. GENERAL GEOLOGY The Cuyuna Range is known for its complicated geologic structure, complex orebodies, and generally indifferent response to attempts at conventional beneficiation. Several geologists have made intensive studies of the Cuyuna Range, including Leithl (1907), ~darns' (1910), Zapffe3 (1925), Grout and Wolff4 (1955), and schmidt5 (1963), whose study was the most comprehensive of all. The Range is covered with glacial deposits - till, out-wash, and lake sediments -up to 250 ft thick. Bedrock stratigraphic units of the Cuyuna Range consist of three tightly folded Precambrian sedimentary formations. The lower unit is the Mahnomen formation, composed of 2000 or more feet of argillite, slate, siltstone, and quartzite. The middle unit is the Trommald formation, which varies from 45 to 500 ft in thickness and comprises the major iron-bearing formation. The upper unit is the Rabbit Lake formation, consisting of over 2000 it of argillite, slate, and interbedded lenses of lean iron formation. In the central part of the Cuyuna Range, the units have been folded into a series of

Jan 1, 1969

By E. Yan, M. Mankosa, E. B. Wasmund

Mechanical cells are the dominant unit operation in base metal rougher flotation applications worldwide. As economic demand for metals increase, and as the feasible ore grades decrease, the installed capacity of mechanical flotation units worldwide has greatly increased, adding significant capital and operating cost. One opportunity to improve this situation is to use fundamental knowledge to make the industrial flotation process more efficient. A key insight is that the flotation process can be optimized by dividing the processing of the pulp into two distinct fluidic environments; a collection cell where air and pulp mix in a highly energetic compartment, and a low energy gravity separation cell. This is the basis of the “2-stage flotation device” that has been patented by Eriez and marketed under the tradename StackCell™. Two industrial case studies will be presented to highlight the potential metallurgical and commercial advantages of the StackCell. In the first, a train consisting of three pilot scale StackCells (each 0.6 metre diameter) was run in parallel with a train of conventional mechanical cells in a major copper concentrator and benchmarked against a batch Denver test. In the second, the performance of a 3 metre diameter StackCell in a nickel sulfide cleaning application was benchmarked against a batch Denver test. The flotation kinetics observed in the StackCells were 2.4 to 2.9 times faster when compared with the lab Denver test. In the first case study, the kinetics of the StackCell was about six times faster than conventional mechanical cells. It is hypothesized that the efficiency improvement in kinetics is because of reduced “drop-back” during the froth recovery phase, which is reduced because of the reduction in the shear that is present in a conventional mechanical tank cell. This result suggests that a 2-stage flotation system can be used to reduce the working volume of flotation units by five to six times. This would allow operators to significantly reduce the size of their flotation lines for a comparable flotation objective. Some discussion will follow about the scale-up criteria, maintenance and operability issues, comparisons of layout, and possible cost savings. INTRODUCTION The solution that most equipment suppliers have offered to reduce the cost of mechanical flotation equipment over the last 30 years has been to increase the size of the flotation unit cell (Mankosa, 2017). For example, three 100 m3 machines with three smaller motors, mechanisms and control valves could be replaced with a geometrically similar 300 m3 with a larger version of each ancillary component. While there are some efficiencies and cost improvements to be gained by this approach, it does not take advantage of fundamental insights into the flotation process, and it does not represent a step-change in terms of process design or economic benefit for the customer. The advantage of the “bigger is better” approach is that there is past experience of incremental improvements by making cells larger and there is a well-established scale-up method, which allows practitioners to estimate equipment capital costs based on very simple laboratory test- work. A different paradigm, which is gaining interest rapidly is to use a unit operation in which it is easier to optimize the flotation process, which will be described in this paper. This approach is often referred to “2- stage flotation”. Because this approach is new, it is important to benchmark it against existing conventional technology and to demonstrate a robust methodology for sizing equipment and flowsheets. This paper aims to address these issues with two case studies.

Jan 1, 2019

By Parer W. R

During the last fifteen years there has been a marked change in the demand and supply patterns of coking coal. The world demand, predominantly Japan, has increased from 23 million tonnes of world export/import coal in 1960 to an expected total of at least 100 million tonnes by 1975. In the last ten years Australia has been the main beneficiary in the growth of the export coal trade. This has been brought about by its proximity to Japan and its large econ- omically mineable reserves of coking coal. A feature of the recent coal development in Australia has been the requirement of companies in new areas to provide massive capital for infrastructure costs. The ability of the Australian coal industry to retain a large share of the world export market depends to a large extent on reasonable costs to transport coal from mine to the market place. In some areas of Australia this means that railway lines, loading and unloading facilities, and ports will have to be up-dated at substantial costs. On the long haul to Europe the recent increases in the prices of bunker fuel have made the price of coal less competitive when compared with sources

Jan 1, 1974

By John A. L. Campbell

Although there has been an increased interest in the electrokinetic properties of coal (1) research on the subject is still in an early stage of development (2,3,4,5,6,7) To the author's knowledge, no electrokinetic studies have been made on coal lithotypes - particularly anthracite lithotypes. This paper presents the results of an investigation of the double layer properties displayed at the anthracite/water interface (whole coal and lithotypes) as determined by streaming potential methods, It is hoped that this investigation and related future investigations will provide correlations from which the surface properties of any coal can be predicted and then related to surface dependent processes such as froth flotation and flocculation. Another aspect of this research, undertaken to determine the physical and chemical parameters necessary for anthracite lithotype differentiation, was previously published (8). The number of parameters involved, lithologic, organic and inorganic, and the wide divergence of these parameters within an individual coal limits the practical utilization of the results, however, the results are of value and of interest in guiding theoretical developments. Keeping the limitations in mind, the results are analyzed in terms of their application to froth flotation processes and are used to indicate the feasibility of making a lithotype separation by flotation.

Jan 1, 1969

By Calvin J. Konya

Precision Blasting has designed easy to use, yet, highly powerful databases for storage and retrieval of blasting data. Three different Databases are available for blasting data management. The Vibration Data Management is a complete vibration record keeping system. The Blasting Data Management will record the information from the blasting report as well as the vibration data. The Explosives Inventory Control makes the control of explosives and initiators quick and easy.

Jan 1, 1991

By A. S. Peck

This report presents the results of a cooperative study between the Bureau of Mines and the University of Utah to define the mechanism involved in the adsorption of oleic acid and sodium oleate on fluorite, calcite, and barite, using an infrared spectrophotometer for identifying and measuring the adsorbed reagents and reaction products. Fluorite, calcite, and barite react with oleic acid and sodium oleate in a water system, forming chemisorbed and solution products. The chemisorbed species is a surface mineral cation-collector anion salt, Chemical analysis of solution resulting from the adsorption reaction indicates that the adsorption process causes displacement of mineral anions, namely, fluoride, carbon ate, and sulfate from the surface of these minerals.

Jan 1, 1963

By Boyling A

United Collieries Pty Limited is an Australian mining company operating one underground coal mine, United Colliery at Warkworth in the Hunter Valley district of New South Wales. Centrally positioned, United Colliery is adjoined by other coal mining operations and is located 80km north - west of Newcastle and 20 kilometres south - west of Singleton. (Fig. 1) High standard road and rail links provide ready access to customers and suppliers.

Jan 1, 1998

By F. P. Haver

The Bureau of Mines developed a process to extract about 90 percent of the sulfur from molybdenite flotation concentrate by compacting the concentrate with one-quarter of its weight of aluminum powder, heating it at 800° C for 30 minutes, and then allowing the product to react with water. This thermite reduction yields an impure molybdenum alloy plus aluminum sulfide. The latter readily hydrolyzes, to yield pure hydrogen sulfide which can be converted to elemental sulfur by the Claus process. At least 95 percent of the molybdenum can be subsequently recovered as 99.95-percent-pure molybdic oxide.

Jan 1, 1968

By Gobbs C., J. Lam

"The Eagle East project, accessed from the existing and operating Eagle Mine in the Upper Peninsula of Michigan, is an underground base metal project currently in execution by Lundin. This paper will provide an update on several aspects of ventilation including use of a twin-decline access, innovative airflow crossover arrangement and construction ventilation methodology. Integration into existing mine ventilation systems including construction methodology will also be covered. Performance benchmarks and lessons learned are discussed across these topics.INTRODUCTIONLundin has owned and operated the Eagle Mine since 2013. This hard rock underground mine is located in the upper peninsula of Michigan, USA. High-grade nickel ore is mined using sublevel open stoping mining methods and trucked to surface via haul ramp. The mine, having been in production for four years, is nearing the completion of the initial orebody. A satellite orebody 2,000 meters East of the current workings and approximately 900 meters below surface, was identified that could extend operations until 2023. In 2016, Lundin approved the Eagle East project and construction began in 2017 (Lundin Mining Corporation, 2017).Due to proximity to the existing mine, Eagle East will be both accessed and serviced from the existing Eagle Mine. Ventilation at the Eagle Mine uses two fresh air intakes with direct fired combustion heaters; one at the mine portal feeding fresh air into the ramp and one at a fresh air raise located near the Eagle deposit. The primary ventilation circuit is powered on the exhaust air side, with two 700 hp exhaust fans located at a single return air raise. Underground, fresh air is distributed through the main access decline and fresh air raise (FAR) to the production levels and controlled with mechanical regulators on the exhaust side. The exhaust air is collected through return air raises with a collector/transfer drift on 250 Level and then to the surface through a return air raise (RAR)."

Jan 1, 2019

By C. C. Martin, M. C. Fuerstenau, R. B. Bhappu

Experiments revealed that quartz could not be floated in conductivity water at any pH with a long-chained sulfonate as collector. Various cations, Fe+++, Al+++, Pb++, Mn++, Mg++, Ca++, are shown to function as activators when the pH is such that the cation hydrolyzes. As the solubility product of the various cation-sulfonates was exceeded at the concentrations involved, precipitated cation-hydroxy-sulfonates, e.g. Fe (RSO3)2OH must be functioning as the collector in these systems. A number of interesting and important phenomena were revealed during a study of the response of beryl to sulfonate flotation.' Perhaps the most important observation is the effect of various cations on flotation response. Experiments showed that in the vicinity of pH 3.0, about three ppm Fe+++ increased recovery of leached beryl from about 60 pct to 100 pct, whereas about four ppm Ca+ + reduced the recovery from 60 pct to 2 pct. With the exception of ferric iron, all of the other cations investigated reduced flotation recoveries drastically at this pH. As postulated in the paper Sulfonate Flotation of Beryl,' even with low concentrations of polyvalent cations and sulfonate (on the order of one ppm), cation-sulfonates are precipitated immediately in solution. That is, blueish-white clouds were noted to form after given additions of salts to known solutions of sulfonate. Furthermore, it was also shown that one-hundred-fold more ferric iron than necessary to form the precipitated cloud had to be added to effect complete flotation. In other words, some form of solid ferric sulfonate was found to function as the collector in these systems. The form of the compound was suggested to be Fe(RCO3)2OH and that bonding to the surface occurs through the hydrogen of the hy-droxyl. As the form of this compound was shown to be dependent on the extent of hydrolysis of ferric iron, it is to be expected that ferric iron will be unique in some pH region. Further, if the presence of hydroxyl in the neutral precipitated compound is of primary importance, then other cations should function similarly but at different values of pH, since they will hydrolyze at different concentrations of hydroxyl ion. To determine whether the premise of hydrolysis is correct, the study was extended to quartz systems in the presence of various cations at various values of pH. EXPERIMENTAL MATERIALS AND METHOD Sodium alkyl aryl sulfonate was chosen as the collector for the experimental work. The reagent, sup plied as a courtesy sample by the Shell Chemical Co., has the following physical properties: Physical form Solid — Finely ground Sulfonate content, pct wt 95 to 97 Molecular Weight 450 to 470 Number of carbon atoms in hydrocarbon chain 25 to 30 Melting Point, OF 250 to 260 Solubility Soluble in water with gel formation at concentrations higher than 25 pct by weight sulfonate This particular reagent was chosen because of its high active sulfonate content, long-chain length, and solid form. With the exception of the frother and sulfonate, all other chemicals were of reagent grade quality. Conductivity water was used in the experimental work. This water, made by passing distilled water through an ion exchange column, had an average measured conductivity of one micromho. Experiments were conducted at room temperature in a small glass flotation cell with the following procedure: 1) A redetermined amount of water and salt solution (e.g. CaC12 . 2H2O) were combined and the pH adjusted to a given value. 2) A given amount of sulfonate was added so that final solution volume was 130 cc. 3) One drop of Dow 250 frother was added. 4) Five g of quartz were added and the system conditioned for five minutes. 5) The pH of the system was measured (termed flotation pH).

Jan 1, 1963

By Richard W. Lodge

In the crucible-method of assaying ores for silver a certain amount of litharge is essential to supply sufficient lead to collect the precious metals. The object of this paper is to point out that the use of a large excess of' litharge in the assay of some ores will give results for silver that are uneven as well as low. So far as I know, however, an excess of litharge does not affect the results obtained in the crucible-assay of ores for gold. The main reasons for using, in the crucible-assay, much more litharge than is required to givc the necessary lead button are: 1, its action as a flux; 2, its action as a desulphur-izer; and 3, its action as an oxidizer, especially on metals like copper and nickel, whereby they are forced into the slag as oxides and thus prevented from passing into the lead button. Hence, by the use of much litharge in the crucible-assay more ore call often be taken than in the scorification-process, and a lead button obtained which call possibly be cupelled at once or after a single scorification. The method is specially advantageous with an ore carrying much copper or similar impurity and poor in silver, when the assayer does not wish to resort to a wet-analysis for the determination of the silver. It is well known that some ores give better results when assayed by the scorification-method, while others give better results by the crucible-method. For several years I have noticed that, when much litharge was used with certain sulphide and arsenical ores, the results were considerably lower than when the scarification-method was used. This fact was more forcibly brought out in connection with certain work carried on this year by Messrs. H. A. Frame and

Jan 1, 1908