本文選題：銅陽極泥 + 鉛�。� 參考：《太原理工大學(xué)》2017年碩士論文

【摘要】：銅陽極泥是提取貴金屬和稀有金屬的主要原料,對其進行高效綜合回收利用顯得尤為重要。傳統(tǒng)火法工藝處理銅陽極泥會導(dǎo)致部分稀貴金屬的損失,而且產(chǎn)生有毒氣體,污染環(huán)境,危害人們身體健康,因此,如何高效無污染處理銅陽極泥成為亟待解決的問題。本文以銅陽極泥為原料,采用鹽酸、氯化鈉溶液的一種或兩種脫除銅陽極泥中鉛,并考察脫鉛對金提取的影響;通過焦亞硫酸鈉、SO2和水合肼復(fù)合還原劑還原氯化分金后液,得鉑鈀精礦,利用控制電位法脫除鉑鈀精礦中賤金屬并提取金,實現(xiàn)了高效無污染的提取陽極泥中金。研究了氯化體系中鉛的存在形態(tài),并從理論上分析復(fù)合還原氯化分金后液動力學(xué)及熱力學(xué)。采用鹽酸體系、氯化鈉和鹽酸混合體系及氯化鈉體系脫除脫銅陽極泥中鉛。鹽酸體系中,適宜實驗條件是鹽酸質(zhì)量濃度為30%,液固比為6:1,反應(yīng)時間為2 h,反應(yīng)溫度為55℃。氯化鈉和鹽酸體系中,適宜實驗條件是反應(yīng)溫度為55℃,反應(yīng)時間為30 min,氯化鈉用量為200 g/L,液固比為12:1,鹽酸質(zhì)量濃度為12%。氯化鈉體系中,適宜實驗條件是反應(yīng)溫度為55℃,氯化鈉用量為340 g/L,反應(yīng)時間為30 min,液固比為8:1。對氯化體系中鉛的存在形態(tài)進行分析,研究表明,三種體系中,可溶性Pb(Ⅱ)均主要以Pb Cl42-絡(luò)合離子形態(tài)存在。脫銅陽極泥在氯化鈉體系適宜條件下進行二段脫鉛,鉛的浸出率為96.90%,渣率為48.40%,脫鉛后陽極泥中貴金屬得到高度富集,將富集后陽極泥進行氯化分金,金的浸出率為99.80%,未脫鉛陽極泥直接進行氯化分金,金的浸出率為95.20%。以沉金后液及氯化分金液的混合溶液為原料,采用焦亞硫酸鈉、so2和水合肼復(fù)合還原稀貴金屬。正交實驗得出影響硒碲還原率的大小順序依次為反應(yīng)時間、水合肼用量、焦亞硫酸鈉用量。還原硒的適宜實驗條件是反應(yīng)時間為4h,焦亞硫酸鈉用量為15g/l,水合肼用量為2ml/l,還原te的適宜實驗條件是反應(yīng)時間為4h,焦亞硫酸鈉用量為30g/l,水合肼用量為3ml/l,還原渣主要物相成分為單質(zhì)碲,其形貌為顆粒集聚體。焦亞硫酸鈉復(fù)合還原硒碲的反應(yīng)過程符合一級反應(yīng)動力學(xué)規(guī)律,硒、碲的表觀活化能分別為ese=52.53kj/mol,ete=70.83kj/mol,硒碲還原過程屬于化學(xué)反應(yīng)控制,熱力學(xué)分析表明,混合溶液中金鉑鈀硒碲分別為aucl4-、ptcl42-、pdcl42-、h2seo3、tecl62-形態(tài)存在,焦亞硫酸鈉、so2主要以h2so3形態(tài)存在,水合肼以n2h5+形態(tài)存在。以焦亞硫酸鈉復(fù)合還原所得還原渣及鉑鈀精礦的混合渣為原料,研究控制電位法脫除賤金屬并提取金的工藝�？刂齐娢宦然s的適宜條件是液固比為5:1,電極電位為500mv,反應(yīng)時間為3h,鹽酸摩爾濃度為5mol/l,反應(yīng)溫度為90℃,在該條件下進行鉑鈀精礦控制電位除雜,碲的脫除率為43.00%,銅的脫除率為82.10%,金、鉑、鈀不被浸出。經(jīng)控制電位除雜后的鉑鈀精礦,通過na2so3浸出除硒,得貴金屬高度富集的鉑鈀精礦。將除雜后的鉑鈀精礦電極電位控制在900mv下溶解,得含金溶液,通過控制電位法還原金,該工藝適宜的條件是以50g/l的na2s2o5溶液為還原劑,反應(yīng)時間為2 h,電極電位為600 mV,反應(yīng)溫度為60℃。在該條件下進行放大實驗,金的還原率為97.00%,還原渣中金的含量為80.00%,鉑鈀硒碲銅的含量分別為2.53%、0.05%、0.03%、1.26%、0.06%。
[Abstract]:Copper anode slime is the main raw material for the extraction of precious metals and rare metals. It is very important to carry out efficient and comprehensive recovery and utilization of copper anode. The traditional process of processing copper anode slime by traditional fire process will lead to the loss of some rare precious metals, and produce toxic gases, pollute the environment and harm the health of the people. Therefore, how to treat the copper anode slime with high efficiency and no pollution is made. It is an urgent problem to be solved. In this paper, copper anode slime is used as raw material, one or two kinds of Sodium Chloride Solution are used to remove lead in copper anode slime, and the effect of lead removal on gold extraction is investigated. Platinum palladium concentrate is obtained by reduction of chlorinated gold by sodium pyrosulphate, SO2 and hydrazine compound reductant, and platinum palladium concentrate is removed by control potential method. The medium metal and gold are extracted and gold is extracted from the anode mud with high efficiency and non pollution. The existence form of lead in the chlorination system is studied. The hydrokinetics and thermodynamics after the compound reduction of the chlorinated gold are theoretically analyzed. The hydrochloric acid system, the mixture of sodium chloride and hydrochloric acid and the sodium chloride system are used to remove the lead in the decopper anode mud. The suitable experimental conditions are the mass concentration of hydrochloric acid 30%, the liquid to solid ratio 6:1, the reaction time 2 h and the reaction temperature 55. The suitable experimental conditions are the reaction temperature 55, the reaction time 30 min, the dosage of sodium chloride 200 g/L, the liquid to solid ratio 12: 1, the mass concentration of hydrochloric acid in the 12%. NaCl system. The condition is that the reaction temperature is 55 C, the dosage of sodium chloride is 340 g/L, the reaction time is 30 min, the liquid solid ratio is 8:1., the existence form of lead in the chlorination system is analyzed. The study shows that in the three systems, the soluble Pb (II) is mainly in the form of Pb Cl42- complex ions. The decopper anode slime is delead two segments under the suitable condition of the sodium chloride system. The leaching rate of lead is 96.90% and the slag rate is 48.40%. The high concentration of precious metals in the anode mud after lead removal is highly enriched, and the anode mud is enriched after enrichment. The leaching rate of gold is 99.80%, and the leaching rate of the anode slime is not directly chlorinated. The leaching rate of gold is 95.20%., and the mixed solution of gold after sinking gold and chlorinated gold is used as the raw material, and Jiao Yaliu is used as the raw material. Sodium acid, SO2 and hydrazine are combined to restore rare precious metals. The orthogonal experiment shows that the order of reducing the reduction rate of selenium and tellurium is the reaction time, the dosage of hydrazine hydrate and the dosage of sodium pyrosulfite. The suitable experimental conditions for the reduction of selenium are the reaction time is 4h, the dosage of sodium hyalurite is 15g/l, the dosage of hydrazine is 2ml/l, and the suitable experimental condition for reducing te is the reduction of selenium. The reaction time is 4h, the dosage of sodium pyrosulfite is 30g/l, the dosage of hydrazine hydrate is 3ml/l, the main substance of the reducing slag is divided into single tellurium, and its morphology is particle agglomeration. The reaction process of reducing selenium and tellurium with sodium pyrosulfite conforms to the first order reaction kinetics law. The apparent activation energy of selenium and tellurium is ese=52.53kj/mol, ete=70.83kj/mol, and selenium tellurium, respectively. The original process belongs to chemical reaction control. The thermodynamic analysis shows that the gold, platinum, palladium, selenium and Tellurium in the mixed solution are aucl4-, ptcl42-, pdcl42-, h2seo3, tecl62-, and SO2 mainly exist in H2SO3 form, hydrazine hydrate is in the form of n2h5+, and the mixture residue of the reduction residue and platinum palladium concentrate from the compound reduction of sodium pyrosulfite is the original. The control potential method for removing the base metal and extracting gold is studied. The suitable conditions for controlling the removal of impurity are that the liquid to solid ratio is 5:1, the electrode potential is 500mv, the reaction time is 3h, the molar concentration of hydrochloric acid is 5mol/l, the reaction temperature is 90 C. Under this condition, the control potential removal of the platinum palladium concentrate, the removal rate of tellurium is 43%, and the copper is removed. The rate is 82.10%, gold, platinum, palladium is not leached. Platinum palladium concentrate with high concentration of precious metals is obtained by Na2SO3 leaching by controlling the removal of platinum and palladium concentrate after controlled potential removal. The electrode potential of platinum palladium concentrates after impurity removal is dissolved under 900mv, and gold containing solution is obtained by controlling potential method. The suitable condition of this process is Na2S2O5 of 50g/l. The solution is a reductant, the reaction time is 2 h, the electrode potential is 600 mV, and the reaction temperature is 60. Under this condition, the reduction rate of gold is 97%, the gold content in the reduced slag is 80%, the content of platinum, palladium, selenium and tellurium copper is 2.53%, 0.05%, 0.03%, 1.26%, 0.06, respectively.
【學(xué)位授予單位】：太原理工大學(xué)
【學(xué)位級別】：碩士
【學(xué)位授予年份】：2017
【分類號】：TF831

【參考文獻】

相關(guān)期刊論文 前10條

1 馬亞峗;鄭雅杰;丁光月;王俊文;董俊斐;張福元;;鹵素離子催化作用下SO_2還原沉金后液及其熱力學(xué)特征[J];中國有色金屬學(xué)報;2016年04期

2 金哲男;馬致遠;楊洪英;劉新建;;銅陽極泥全濕法處理過程中貴賤金屬的行為[J];東北大學(xué)學(xué)報(自然科學(xué)版);2015年09期

3 張福元;鄭雅杰;孫召明;馬亞峗;董俊斐;;采用亞硫酸鈉還原法從沉金后液中回收稀貴金屬[J];中國有色金屬學(xué)報;2015年08期

4 李春俠;王愛榮;陳繼平;;銅陽極泥處理半濕法工藝與卡爾多爐工藝的比較[J];企業(yè)技術(shù)開發(fā);2015年19期

5 侯凱;謝賢;童雄;崔毅琦;呂昊子;孟奇;陸婭琳;;我國金礦床的工業(yè)類型及選礦研究方法[J];礦產(chǎn)綜合利用;2014年04期

6 聶華平;張忠堂;王日;黃紹勇;王莉紅;;Au-Cl~--SO_3~(2-)-H_2O系的電化學(xué)行為研究[J];濕法冶金;2014年04期

7 焦志良;陳為亮;張旭;朱玉平;許康;齊妍潔;;從二次含鉛物料中濕法回收鉛的研究現(xiàn)狀[J];濕法冶金;2014年02期

8 簡椿林;;某復(fù)雜金精礦的金銀提取工藝研究[J];黃金;2014年02期

9 張福元;卓儉進;張玉明;彭國敏;;沉淀轉(zhuǎn)化法回收酸浸渣中鉛的酸溶工藝研究[J];無機鹽工業(yè);2013年11期

10 阮書鋒;尹飛;王成彥;王軍;王振文;;H_2SO_4+NaCl選擇性浸出鉛陽極泥的研究[J];礦冶;2012年03期

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